Saimm 201508 aug

VOLUME 115

NO. 8

AUGUST 2015

Despite current challenges facing the min ning and minerals indusstry, Level 1 BEE Contributor, MIP Process Technologies (“MIP�) continues to pursue a positive standing in itss operations.

According to MD, Philip Hoff, they have enhanced its staff base to t focus on its business areas, both locally and abroad. “We have recently ap ppointed ho will be a General Ma anager, Rogan Roulstone, wh responsible fo or the day-to-day operations. Rogan comes with vast experience in project, pro ocess and general mana agement�. “To date our success s can be attributed to o upholding customer focus, reliable products and con ntinuous astructure improvementt and development. This infra will continue tto be the back bone of our company,� c says Hoff, “as we believe that we have exxcelled ugh our in customer se ervice and good value throu cessing knowledge. onjunction minerals proc knowledge This, This in co with engineerring expertise enables us to a adapt our designs to suitt each and every client’s nee eds.�

Since inceptio on in 2007, MIP’s endeavourss onto the international stage s have been very substa antial. They have establish uccessfully hed equipment operating su in Turkey, Australia, North America, Canad da, and

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! destinations. y has actively pursued the exxport The company market, as Ho an learn a off believes that the world ca vast amount from f local technology. “Soutth African expertise in th he minerals and mining industry is highly regarded thro oughout the world and we do have the products to prove it. We can boast a number of ! #

h, is the largest Linear Screenss in the One of which world and the ese were followed up with subsequent orders,� conc cludes Hoff.

MIP designs a and supplies a range of world d class process equip pment which includes Thickeners,

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" Samplers and d Linear Screens. It has an added string to its bow with h the ADCS range of Dust Exttraction

! renowned Ch hemineer range of Mixers and d Agitators.

“MIP offers the com mplete ran nge of process and dust exttraction equipme ent – you ur one stop equipment sup pplier� conclude es Hoff.

The Southern African Institute of Mining and Metallurgy OFFICE BEARERS AND COUNCIL FOR THE 2014/2015 SESSION Honorary President Mike Teke President, Chamber of Mines of South Africa Honorary Vice-Presidents Ngoako Ramathlodi Minister of Mineral Resources, South Africa Rob Davies Minister of Trade and Industry, South Africa Naledi Pandor Minister of Science and Technology, South Africa President J.L. Porter President Elect R.T. Jones Vice-Presidents C. Musingwini S. Ndlovu Immediate Past President M. Dworzanowski Honorary Treasurer C. Musingwini Ordinary Members on Council V.G. Duke M.F. Handley A.S. Macfarlane M. Motuku M. Mthenjane D.D. Munro G. Njowa

T. Pegram S. Rupprecht N. Searle A.G. Smith M.H. Solomon D. Tudor D.J. van Niekerk

Past Presidents Serving on Council N.A. Barcza R.D. Beck J.A. Cruise J.R. Dixon F.M.G. Egerton G.V.R. Landman R.P. Mohring

J.C. Ngoma S.J. Ramokgopa M.H. Rogers G.L. Smith J.N. van der Merwe W.H. van Niekerk

Branch Chairmen Botswana

L.E. Dimbungu

DRC

S. Maleba

Johannesburg

I. Ashmole

Namibia

N.M. Namate

Northern Cape

C.A. van Wyk

Pretoria

N. Naude

Western Cape

C. Dorfling

Zambia

D. Muma

Zimbabwe

S. Ndiyamba

Zululand

C.W. Mienie

Corresponding Members of Council Australia: I.J. Corrans, R.J. Dippenaar, A. Croll, C. Workman-Davies Austria: H. Wagner Botswana: S.D. Williams United Kingdom: J.J.L. Cilliers, N.A. Barcza USA: J-M.M. Rendu, P.C. Pistorius

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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)

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H. Britten (1955–1956) Wm. Bleloch (1956–1957) H. Simon (1957–1958) M. Barcza (1958–1959) R.J. Adamson (1959–1960) W.S. Findlay (1960–1961) D.G. Maxwell (1961–1962) J. de V. Lambrechts (1962–1963) J.F. Reid (1963–1964) D.M. Jamieson (1964–1965) H.E. Cross (1965–1966) D. Gordon Jones (1966–1967) P. Lambooy (1967–1968) R.C.J. Goode (1968–1969) J.K.E. Douglas (1969–1970) V.C. Robinson (1970–1971) D.D. Howat (1971–1972) J.P. Hugo (1972–1973) P.W.J. van Rensburg (1973–1974) R.P. Plewman (1974–1975) R.E. Robinson (1975–1976) M.D.G. Salamon (1976–1977) P.A. Von Wielligh (1977–1978) M.G. Atmore (1978–1979) D.A. Viljoen (1979–1980) P.R. Jochens (1980–1981) G.Y. Nisbet (1981–1982) A.N. Brown (1982–1983) R.P. King (1983–1984) J.D. Austin (1984–1985) H.E. James (1985–1986) H. Wagner (1986–1987) B.C. Alberts (1987–1988) C.E. Fivaz (1988–1989) O.K.H. Steffen (1989–1990) H.G. Mosenthal (1990–1991) R.D. Beck (1991–1992) J.P. Hoffman (1992–1993) H. Scott-Russell (1993–1994) J.A. Cruise (1994–1995) D.A.J. Ross-Watt (1995–1996) N.A. Barcza (1996–1997) R.P. Mohring (1997–1998) J.R. Dixon (1998–1999) M.H. Rogers (1999–2000) L.A. Cramer (2000–2001) A.A.B. Douglas (2001–2002) S.J. Ramokgopa (2002-2003) T.R. Stacey (2003–2004) F.M.G. Egerton (2004–2005) W.H. van Niekerk (2005–2006) R.P.H. Willis (2006–2007) R.G.B. Pickering (2007–2008) A.M. Garbers-Craig (2008–2009) J.C. Ngoma (2009–2010) G.V.R. Landman (2010–2011) J.N. van der Merwe (2011–2012) G.L. Smith (2012–2013) M. Dworzanowski (2013–2014)

Honorary Legal Advisers Van Hulsteyns Attorneys Auditors Messrs R.H. Kitching Secretaries The Southern African Institute of Mining and Metallurgy Fifth Floor, Chamber of Mines Building 5 Hollard Street, Johannesburg 2001 P.O. Box 61127, Marshalltown 2107 Telephone (011) 834-1273/7 Fax (011) 838-5923 or (011) 833-8156 E-mail: journal@saimm.co.za

The Journal of The Southern African Institute of Mining and Metallurgy

Editorial Board

Editorial Consultant D. Tudor

Typeset and Published by The Southern African Institute of Mining and Metallurgy P.O. Box 61127 Marshalltown 2107 Telephone (011) 834-1273/7 Fax (011) 838-5923 E-mail: journal@saimm.co.za

Printed by Camera Press, Johannesburg

Advertising Representative Barbara Spence Avenue Advertising Telephone (011) 463-7940 E-mail: barbara@avenue.co.za The Secretariat The Southern African Institute of Mining and Metallurgy ISSN 2225-6253 (print) ISSN 2411-9717 (online)

THE INSTITUTE, AS A BODY, IS NOT RESPONSIBLE FOR THE STATEMENTS AND OPINIONS ADVANCED IN ANY OF ITS PUBLICATIONS. Copyright© 1978 by The Southern African Institute of Mining and Metallurgy. All rights reserved. Multiple copying of the contents of this publication or parts thereof without permission is in breach of copyright, but permission is hereby given for the copying of titles and abstracts of papers and names of authors. Permission to copy illustrations and short extracts from the text of individual contributions is usually given upon written application to the Institute, provided that the source (and where appropriate, the copyright) is acknowledged. Apart from any fair dealing for the purposes of review or criticism under The Copyright Act no. 98, 1978, Section 12, of the Republic of South Africa, a single copy of an article may be supplied by a library for the purposes of research or private study. No part of this publication may be reproduced, stored in a retrieval system, or transmitted in any form or by any means without the prior permission of the publishers. Multiple copying of the contents of the publication without permission is always illegal.

U.S. Copyright Law applicable to users In the U.S.A. The appearance of the statement of copyright at the bottom of the first page of an article appearing in this journal indicates that the copyright holder consents to the making of copies of the article for personal or internal use. This consent is given on condition that the copier pays the stated fee for each copy of a paper beyond that permitted by Section 107 or 108 of the U.S. Copyright Law. The fee is to be paid through the Copyright Clearance Center, Inc., Operations Center, P.O. Box 765, Schenectady, New York 12301, U.S.A. This consent does not extend to other kinds of copying, such as copying for general distribution, for advertising or promotional purposes, for creating new collective works, or for resale.

VOLUME 115

NO. 8

AUGUST 2015

Contents Journal Comment, by R.E. Robinson . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

President’s Corner, by J.L. Porter . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Career Development in the Minerals Industry Event: SAIMM Young Professionals Council

by V. Maseko . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

iv–v vii vi

The physical ability of women in mining: can they show muscle?

by D. Botha, and J.F. Cronjé . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Transformation in the South African mining industry – looking beyond the employment equity scorecard

by N.V. Moraka and M. Jansen van Rensburg. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Reconciliation along the mining value chain

by A.S. Macfarlane . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Effects on entrainment of serpentines by hydrophobic flocs of ultra-fine copper-nickel sulphides during flotation

by M. Tang, S. Wen, and X. Tong . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Smelting of calcined basic nickel carbonate concentrate in a 200 kW DC arc furnace

by M. Abdellatif . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Comparison of physical properties of oxidative sintered pellets produced with UG2 or metallurgical-grade South African chromite: a case study

by R.I. Glastonbury, J.P. Beukes, P.G van Zyl, L.N. Sadiki, A. Jordaan*, Q.P. Campbell, H.M. Stewart, and N.F. Dawson . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

The influence of selected biomass additions on the co-pyrolysis with an inertinite-rich medium rank C grade South African coal

by C.A. Strydom, T.Z. Sehume, J.R. Bunt, and J.C. van Dyk . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Mine Occupational Safety and Health Leading Practice Adoption System (MOSH) examined – the promise and pitfalls of this employer-led initiative to improve health and safety in South African Mines

by M. Hermanus, N. Coulson, and N. Pillay. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Application of the attainable region technique to the analysis of a full-scale mill in open circuit

by F.K. Mulenga and M.M. Bwalya . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Thermogravimetric investigation of macadamia nut shell, coal, and anthracite in different combustion atmospheres

by S.O. Bada, R.M.S. Falcon, L.M. Falcon, and M.J. Makhula . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Application of systemic flow-based principles in mining

by J.O. Claassen. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Evaluating the coal bump potential for gateroad design in multiple-seam longwall mining: a case study

by X. Wanga, J. Baia, W. Lia, B. Chena, and V.D. Daob. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Influence of medium particle size on the separation performance of an air dense medium fluidized bed separator for coal cleaning

by S. Mohanta and B.C. Meikap . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Measuring and modelling of density for selected CaO-Al2O3-MgO slags

by J.F. Xu, K. Wan, J.Y. Zhang, Y. Chen, and M.Q. Sheng . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Investigation of factors influencing blending efficiency on circular stockpiles through modelling and simulation

by Z. Loubser and J de Korte . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Investigation into strata behaviour and fractured zone height in a high-seam longwall coal mine

by G. Song and S. Yang. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Parametric estimation of capital costs for establishing a coal mine: South Africa case study

by M. Mohutsiwa and C. Musingwini . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

659

669 679

687 691

699

707

717 729

741 747

755

761 767

773

781 789

International Advisory Board VOLUME 115

NO. 8

AUGUST 2015

R. Dimitrakopoulos, McGill University, Canada D. Dreisinger, University of British Columbia, Canada E. Esterhuizen, NIOSH Research Organization, USA H. Mitri, McGill University, Canada M.J. Nicol, Murdoch University, Australia E. Topal, Curtin University, Australia

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AUGUST 2015

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R.D. Beck J. Beukes P. den Hoed M. Dworzanowski B. Genc M.F. Handley R.T. Jones W.C. Joughin J.A. Luckmann C. Musingwini J.H. Potgieter R.E. Robinson T.R. Stacey

Journal Comment Platinum in the 21st Century ‘There is a tide in the affairs of men’ Shakespeare

The theme of most of my previous Journal Comments has been job creation. This demands identifying the goods and services for sale, and its corollary, the skills and education required.

On this latter score, the mining professional community has had some spectacular success in the 21st century. Many hundreds of skilled mining engineering graduates and diplomates have been produced using the latest in computer-based education, a good percentage being women.

In the last few months there have been some remarkable announcements regarding the hydrogen fuel cell – the platinum content alone would probably be equal in quantity to the total annual gold production at the peak period of the last ‘golden’ 20th century.

Since a number of the papers in this issue of the Journal can be related to platinum, I started planning my Comment around this topic. I soon realized that the mining industry might be facing the possibility of replicating the successes of the previous century. I had visions of a suite of mining operations around the full perimeter of the Bushveld Complex. I termed these ‘mining clusters’ to signify their potential to attract a host of other job-creating options and spark a revolution in education at all levels, thanks to digital information technology.

As regards the first paper: ‘Transformation in the South African mining industry – looking beyond the employment equity scorecard ’ by N.V. Moraka and M. Jansen van Rensburg.

Job creation is more productive than job dictation.

Could such discussions move away from the legalistic semantics of BEE scorecards into a national commitment to job creation? Could the whole mining industry not invoke the cooperation of Mintek, the CSIR, IDC, the departments of Minerals and Energy, Trade and Industry, Water Affairs, and the agricultural industry to become stakeholders and participants in the innovative development portfolios about to be undertaken by mining and metallurgical, industrial, and agricultural clusters? The target is an affluent first-world nation in this 21st century.

From lofty ideals to nuts and bolts in the remaining papers in this issue …

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There are many job creation opportunities indicated in the assignments by A.S. Macfarlane in the paper ‘Reconciliation along the mining value chain’, which can be coupled with the paper ‘The physical ability of women in mining: can they show muscle ?’ by D. Botha and J.F. Cronjé’ I gather Macfarlane aims to establish materials balances at every stage of mining and processing, i.e. the added value links in the chain; to be reported in the mine call factors (MCFs). An enormous undertaking, particularly if codes of practice are envisaged.

In the mid 1950s, when I was deeply involved in investigating gold losses, the MCF was legally required to be reported monthly to the police via the Government Mining Engineer in order to combat theft of gold from the plants. (Large thefts have been reported recently.) But, as demonstrated recently in a paper to this Journal, the methods used for sampling a stope face are most unreliable. Since this first link is suspect, the usefulness of the subsequent links in the chain is undermined. Perhaps the contempt shown towards MCFs by gold and platinum mine personnel is well justified.

The eighth paper is: ‘Mine Occupational Safety and Health Leading Practice Adoption System (MOSH) examined – the promise and pitfalls of this employer-led initiative to improve health and safety in South African Mines ’ by M. Hermanus, N. Coulson, and N. Pillay It represents a report-back by a group established at the University of the Witwatersrand, representing the employees known as MOSH. They focus on outcomes and activities on the health and safety aspects of mining. The most important outcome is that we are moving forward to match the statistics of leading international companies.

Most interesting, and hopefully welcome as the harbinger of more to come, is a contribution from China: ‘Effects on entrainment of serpentines by hydrophobic flocs of ultra-fine copper-nickel sulphides during flotation ’ by M. Tang, S. Wen, and X. Tong.

At first sight this paper, treating ore from the Yumang deposit in China, has little relevance to platinum. However, it will be noted that the orebody

The Journal of The Southern African Institute of Mining and Metallurgy

Journal Comment (continued)

The Journal of The Southern African Institute of Mining and Metallurgy

The ninth paper in this selection, and possibly the most erudite in terms of fundamental work on mineral processing and highly significant in the smelting route for platinum recovery from the UG2 reefs, is ‘Application of the attainable region technique to the analysis of a full-scale mill in open circuit’ by F.K. Mulenga and M.M. Bwayla The attainable region technique is one of the great advances in theoretical mathematical modelling algorithms, developed by David Glasser and Diana Hildebrand (incidentally one of the notable female chemical engineers) from the University of the Witwatersrand, which is the home of metallurgical research.

Predicting the performance of open circuit ball milling is possibly a forerunner of many more advances in the mathematical modelling of mineral (and even mining) technology. The correlations between the model and experimental results are promising. This work may be a step forward in circumventing the comminution problems associated with the handling of the UG2 reefs.

In the paper, ‘Comparison of physical properties of oxidative sintered pellets produced with UG2 or metallurgical grade South African chromite: a case study ’, by R.I. Glastonbury, J.P. Beukes, P.G. van Zyl, L.N. Sadiki, A. Jordaan, Q.P. Campbell, H.M. Stewart, and N.F. Dawson meticulously demonstrate that the chromites obtained from the final residues from the processing of the UG2 Reef for platinum group metals are every bit as favourable for manufacturing sintered pellets for ferrochrome production as the standard material. This material could be of value to both the platinum and the ferrochrome producers in terms of its superior properties and mutually advantageous prices. There are many other options for using the platinum fuel cell and for job creation in conjunction with South Africa’s wealth of other mineral resources

R.E. Robinson

AUGUST 2015

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contains approximately 2 ounces of platinum and palladium per ton. This shows that the platinum metals are more common than I thought! The base metals in the ore are much the same as those in the BC ores. Availability of water, and the recycling and disposal of aqueous effluents are common considerations in hydrometallurgy. If expansion takes place in the platinum sector, the opportunity for research, particularly on the use of the platinum fuel cell, will become attractive and international cooperation could be worthwhile. Since Chinese investors are already involved in South Africa, it is worth suggesting they may become stakeholders in the research portfolios.

Basic nickel carbonate is a material that can be readily available from the proposed hydrometallurgical ‘Kell’ process for platinum group metals recovery. The paper ‘Smelting of calcined basic nickel carbonate concentrate in a 200 kw DC arc furnace ’ by M. Abdellatif indicates that this material can be smelted to produce nickel metal which could be the preferred form in sales. Moreover, the DC arc furnace could be powered by the hydrogen fuel cells, manufactured locally to produce low cost power.

There are alternative options for sources of hydrogen. One is in the form of ‘syngas’ (a mixture of hydrogen and CO) which can be produced by the pyrolysis of carbonaceous plant material such as bagasse, shells of nuts, and sawdust from timber waste. A South African option named the ‘Beauti-fuel’ process is available among the many other options to form a worthwhile research portfolio.

The pyrolysis of such materials is described in two papers: ‘Thermogravimetric investigation of macadamia nut shell – coal and anthracite in different combustion atmospheres’ by S.O. Bada, R.M.S. Falcon, L.M. Falcon, and M.J. Makhula ‘The influence of selected biomass additions on the co-pyrolysis with an inertinite-rich medium rank C grade South African coal ’ by C.A. Strydom, T.Z. Sehume, J.R. Bunt, and J.C. van Dyk. Since it is possible that the ‘Platinum Province’ will be associated with agricultural activities, particularly the rural small-lot farmers, it is certainly worthwhile to consider such possibilities. The second of the papers introduces the uses of coal materials. There are billions of tons of waste coal discards which can be considered for processing to recover the coal macerals as well as the other valuable constituents in the ash such as iron, sulphur, alumina, and silica. Once again the call for stakeholders in a portfolio of such projects can be usefully considered.

Career Development in the Minerals Industry Event The SAIMM Young Professionals’ Council (SAIMM-YPC)

O

On 31 July 2015, the SAIMM Young Professionals Council (YPC) hosted the Career Development in the Minerals Industry event for final-year university students in mining and metallurgy, at Mintek in Randburg. The event was attended by students from the universities of Johannesburg, Pretoria, and South Africa.

The first Career Development in the Minerals Industry event, organized by the former Career Guidance and Education Committee of the SAIMM, took place in August 2013. This event was not held in 2014. The YPC, having taken over the mandate of promoting the interests of young people in the minerals industry (formerly assigned to the Career Guidance and Education Committee), took on the task of convening the event in 2015, as they considered it valuable for final-year university students.

The event was aimed at providing the students with relevant information as they embark on their professional careers. The motivation was the perception that new graduates entering the minerals industry are often not fully aware of the realities and challenges that they will encounter, which leads to frustration in the early years of their careers.

As the intake of new graduates into the South African mining sector seems to have plateaued, with a significant number of recent graduates finding themselves without work, the issues of employment and employability were high on the agenda of this year’s event. Career and personal development were also recognized as important topics, and featured strongly in the programme.

The programme included representatives from various mining and consulting companies, the Chamber of Mines, the Engineering Council of South Africa, the Engineering Leadership Academy, and the Financial Planning Institute of South Africa, who gave talks on topics such as mentorship, networking, family planning, and financial planning, and which also included some personal testimonies. The event also featured three panel discussions, where questions were put to a panel of industry representatives by the YPC and the students in attendance.

The three panel discussions were on the following topics:

➣ Employment and employability: how to obtain employment and differentiate yourself in the depressed 2015/2016 mining environment ➣ Career development: developing a good career ➣ Personal development: becoming a well-rounded individual.

Based on feedback received from the students, the key learnings that were taken away from the event included (but were not limited to): ➣ South Africa is producing more graduates than the mining industry can handle, The rest of Africa is a growth market for mining, look there for opportunities. Learn languages such as French or Portuguese, to make penetration into the rest of Africa easier ➣ Don’t be too picky with the opportunities that present themselves. The current mining environment doesn’t allow for that.

Take the opportunities that you get and make the best of them ➣ Your university education actually provides you with the tools that you can use to be entrepreneurial. Recognize these tools and use them to be more enterprising ➣ Due to the significant increase in the number of graduates in industry in recent years, the environment has become more competitive. You need to differentiate yourself by going the extra mile and doing what others may not be willing to do ➣ To be successful in industry you need to understand people, be an effective communicator, and show leadership ➣ Don’t underestimate the influence that mentoring and personal networks can have on your professional success.

The YPC has complied a reference handbook for students, which was handed out at the event. The handbook contains information on job search skills, career planning, networking, mentoring, professional registration, and personal development, which students can refer to during the early years of their careers.

The YPC is proud to have organized such an event and we hope that the students benefited considerably from the proceedings of the day. The YPC would also like to thank the many industry participants that contributed to the success of this event, who are unfortunately too many to name individually.

V. Maseko Secretary: Young Professionals’ Council (SAIMM)

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The Journal of The Southern African Institute of Mining and Metallurgy

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ell here it is! This is the last of my ‘President’s Corner’ contributions for my 2014/2015 term of office and, as I commented earlier this year, time moves at an accelerated rate when you are busy. I have thoroughly enjoyed the past year. It has been characterized by the usual highs and lows in terms of our efforts to grow the reach of the Institute geographically as well as building greater relevance to industry and offering value to our members. The latter is critical. After 121 years of successful existence, our Institutional environment is changing faster than ever before. My message here is that we have to be proactive. The status quo is no longer acceptable.

As a result of deep structural changes to the mining industry in Southern Africa in recent years, we have begun to continuously examine how we operate as an Institute and ask ourselves the question: ‘What do we need to do to adapt to changing circumstances?’. For example, several of the larger mining ‘houses’ have reduced the number of staff employed in their corporate offices, with some capacity moving to other geographies; and swings in commodity prices and the ageing of mining assets have curtailed capital expenditure. At the same time, many new junior and mid-tier mining companies have entered our region and we have seen significant shifts in the demographics of our membership. Employment opportunities for our young professionals are a matter of serious concern.

Actions undertaken by the SAIMM to respond to the above, and which you should ALL be well aware of, include the change from ‘South African’ to ‘Southern African ’ in our name; the development of many new Branches throughout the region; the establishment of the Young Professionals Council (YPC); the appointment of a Regional Development role within the SAIMM office; and re-writing our by-laws to bring them in line and to create the required organizational flexibility.

The results so far? Well, we are financially secure for the present. We now have 10 formally established branches; more than 25% of our membership is in the Student category; our YPC is already adding value in terms of its guidance to officebearers and Council. These are all by design, and represent some of the achievements for this and recent years. Is it sufficient?

No, it is NOT sufficient. My comment has to be understood against the backdrop of a depressed global mining industry and some significant regional challenges that impact on the operation of the Institute, as reflected in the overall attendance at our conferences, which has been somewhat of a low point during the year.

In my first contribution to the President’s Corner last year, I closed my article with the following short paragraph: ‘As I have pointed out in my Presidential address, the challenges are severe and the need for action acute. Now is not the time for insular thinking, it is time for greater levels of collaborative thinking and investment than at any time in our mining history. The speed with which we are able to find the collaborative structures and make changes to the mining process in hard rock mines will be the measure of success.’

Mining is an outcome of geological processes, and sometimes I feel as though the much-needed changes in our industry are moving by the geological timescale! I have no qualms in being critical of activities that demonstrate personal and political selfserving interests and which result in this inexplicable inability to move forward, when some of the solutions seem so clear.

Now I am sounding like a sports fan critiquing the team coach from the sideline – nonetheless, many successful coaches listen to their fans.

It is far too easy to fall in to a critical mind-set when the attractiveness of South Africa as a mining destination is dropping in the rankings year after year. When many of our critical indicators of success as a nation are failing in the mining sector. For example, the ability (or inability) to create a stable environment that encourages long-term, risk-based investments; the ability to get national infrastructure in place, not only to create jobs but to create opportunity to grow internationally competitive industries in mining, energy, logistics, and manufacturing; the ability to reduce social conflict and hardship in and around mining communities, etc. These seem to be complex and, at times, intractable issues with few solutions.

However, it is heartening to learn that the Cabinet has given the go-ahead for the Operation Phakisa ‘mining laboratory’ with the objective of identifying problems/issues that are fundamentally constraining growth in the South African mining sector. One would fervently hope that the participants can leave parochial agendas at the door and focus on what is critical to the nation for the sustainable development and exploitation of our natural resources. Anglo CEO Mark Cutifani declares that he will be an ‘enthusiastic participant in the Phakisa Mining Lab’. Can we expect public statements of commitment from other leaders? Are there other multi-stakeholder initiatives that can further the development of concrete actions and not just end in more empty words and reports to be filed away?

As members of the SAIMM it is no longer good enough for us to sit on the sidelines and complain, since we are ALL part of the problem. I invite our members to submit constructive suggestions to the Journal on what YOU believe we should be doing to contribute as a role-player in the mining industry. What can we do to move the issues forward and contribute to activities such as the Mining Laboratory? Is it through the Mining Dialogues 360 initiative, or other vehicles?

We have many highly knowledgeable and experienced members in our ranks who would like to find a way to be engaged.

Let us hear your thoughts, ideas and suggestions …

Finally, for those of you who have taken the time to read my President’s Corner over the past year – thank you for your interest and support.

t’s iden s e r P er Corn

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J.L. Porter President, SAIMM

www.verrefshaped.co.za

Email: sales@verrefshaped.co.za

5 Star Incentive Programme About the SAIMM 5 Star Incentive Programme: The SAIMM are proud to welcome you to our Incentive programme where we have negotiated to provide you with more benefits.

These benefits include: 1. Top 5 Proposers for the current financial year are to be given a free ticket to the SAIMM Annual Banquet with mention at the Annual General Meeting.

2. Top 5 Referees for the financial year are to be given a free ticket to the SAIMM Annual Banquet with mention at the Annual General Meeting.

3. Access to discounts offered by Service providers that have negotiated discounted rates and special offers for you our valued member.

4. Conference attendance within a 2 year period and applies to events that are paid for. If you attend 3 events you get the next conference that you register for at no cost.

5. The author with the most number of papers published in the SAIMM Journal in the previous financial year would be recognised at the Annual General Meeting and will receive a free ticket to the SAIMM Annual Banquet.

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

The physical ability of women in mining: can they show muscle?

by D. Botha, and J.F. Cronjé . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

This paper provides practical recommendations, informed by a literature review and empirical findings, that contribute to the sustainable deployment of women in the mining industry.

Transformation in the South African mining industry – looking beyond the employment equity scorecard

by N.V. Moraka and M. Jansen van Rensburg . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Qualitative insight is provided into the initiatives employed, and challenges experienced, by mining companies in a quest to transform the mining industry. Results from this study show that the assumption that mining companies are reluctant to transform is erroneous, and that there is buy-in and commitment to transformation.

Reconciliation along the mining value chain

by A.S. Macfarlane. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

This paper contends that a full understanding of the metal flow along the value chain, its variability, its underlying loss potential, and its control points is necessary before a systematic approach to reconciliation can be undertaken.

Effects on entrainment of serpentines by hydrophobic flocs of ultra-fine copper-nickel sulphides during flotation

by M. Tang, S. Wen, and X. Tong . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Micro-flotation tests and settling rate tests, as well as visual observations, were used to determine the effects of a combination of strong collectors on entrainment of serpentines in metallic mineral concentrate. The results indicated the presence of serpentines entrapped in the hydrophobic flocs that result from these collectors, even with the use of an effective gangue depressant.

Smelting of calcined basic nickel carbonate concentrate in a 200 kW DC arc furnace

by M. Abdellatif . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Calcined basic nickel carbonate (BNC) concentrate was smelted in a pilot-scale DC arc furnace to produce a metal containing more than 96% Ni. A range of smelting conditions were investigated, with the major variables being reductant type and feed rate, flux composition and addition, and BNC feed rate.

Comparison of physical properties of oxidative sintered pellets produced with UG2 or metallurgical-grade South African chromite: a case study

by R.I. Glastonbury, J.P. Beukes, P.G van Zyl, L.N. Sadiki, A. Jordaan*, Q.P. Campbell, H.M. Stewart, and N.F. Dawson The physical properties of oxidative sintered pellets produced from typical South African UG2 ore and conventional South African metallurgical-grade chromite ore are compared. The results can be utilized by ferrochromium producers to better quantify the advantages and disadvantages associated with the use of UG2 ore for ferrochromium production.

The influence of selected biomass additions on the co-pyrolysis with an inertinite-rich medium rank C South African coal

by C.A. Strydom, T.Z. Sehume, J.R. Bunt, and J.C. van Dyk . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

The co-pyrolysis of four biomass samples (hardwood chip, softwood chip, pinewood chip, and sugarcane bagasse) with an inertinite-rich medium rank C grade South African coal was investigated. The results indicated that the influence of the biomass on the pyrolysis rate of the coal is small and vice versa. The CO2-producing reactions of the coal were slightly enhanced during co-pyrolysis.

Mine Occupational Safety and Health Leading Practice Adoption System (MOSH) examined – the promise and pitfalls of this employer-led initiative to improve health and safety in South African Mines

by M. Hermanus, N. Coulson, and N. Pillay . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

This paper assesses the effectiveness of the Mine Occupational Health and Safety Leading Practice Adoption System (MOSH) and its potential to improve mine health and safety in South African mines. Although the depth of engagement with MOSH amongst stakeholders and on mine sites varied, mining companies, labour representatives, and the Mine Health and Safety Inspectorate (MHSI) saw the programme as significant. continued overleaf

These papers will be available on the SAIMM website

http://www.saimm.co.za

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PAPERS IN THIS EDITION These papers have been refereed and edited according to internationally accepted standards and are accredited for rating purposes by the South African Department of Higher Education and Training Application of the attainable region technique to the analysis of a full-scale mill in open circuit

by F.K. Mulenga and M.M. Bwalya . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

This paper introduces the use of the attainable region (AR) technique to continuous milling. Feed flow rate, ball size, and ball filling were identified as being pivotal for the optimization of open ball milling circuits. Mill speed, on the other hand, had only a limited effect on the production of particles in the size range -75 +10 micron.

Thermogravimetric investigation of macadamia nut shell, coal, and anthracite in different combustion atmospheres

by S.O. Bada, R.M.S. Falcon, L.M. Falcon, and M.J. Makhula. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

The combustion and co-combustion behaviour of macadamia nut shell, high-ash coal, and anthracite, along with their blends, were studied using thermogravimetry. The results provide the combustion and co-combustion characteristics of various samples and their blends and indicate their combustion compatibilities.

Application of systemic flow-based principles in mining

by J.O. Claassen . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Thirty mining operations across the African continent were studied to establish to what extent systemic flow-based principles are applied in day-to-day operations. The results indicate that notable potential exists in most of the operations evaluated to implement flow-based management principles, including geometallurgical principles.

Evaluating the coal bump potential for gateroad design in multiple-seam longwall mining: a case study

by X. Wanga, J. Baia, W. Lia, B. Chena, and V.D. Daob . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

A methodology is proposed for evaluating the risk of coal bumps in multiple-seam longwall mining. By employing the proposed gateroad design, which incorporated a conservative horizontal offset of the longwall panels, coal bumps were avoided completely.

Influence of medium particle size on the separation performance of an air dense medium fluidized bed separator for coal cleaning

by S. Mohanta and B.C. Meikap . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

The medium particle size is shown to have an overriding significance on the separation efficiency in an air dense medium fluidized bed. Experimental results indicate that different size fractions of the same feed coal respond differently to the same size fraction of medium solids, and a particular size fraction of coal responds differently with different size fractions of medium solids.

Measuring and modelling of density for selected CaO-Al2O3-MgO slags

by J.F. Xu, K. Wan, J.Y. Zhang, Y. Chen, and M.Q. Sheng . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

The densities of different slag compositions in the CaO-Al2O3-MgO system were measured at 1823K by the Archimedean method. The results indicate that slag density decreases initially, then increases, with increasing MgO content. A similar trend is seen with increasing CaO/Al2O3 ratio at a fixed MgO content. Application of the molar volume model confirmed that the present expanded approximation rules can be used to predict the molar volumes of the melts tested.

Investigation of factors influencing blending efficiency on circular stockpiles through modelling and simulation

by Z. Loubser and J de Korte . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Simulation models were developed for coneshell and chevcon stacking, and used to investigate the influence of different stockpile parameters on blending efficiency in circular stockpiles.

Investigation into strata behaviour and fractured zone height in a high-seam longwall coal mine

by G. Song and S. Yang . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

A theoretical method is presented to investigate the destabilizing modes of the main roof in a high-seam longwall coal mine (by sliding or rotation), and is used to determine whether the main roof is in the caved or fractured zone.

The fractured zone height is closely related to a number of parameters, with bulking factor, mining height, and roof thickness being the most important.

Parametric estimation of capital costs for establishing a coal mine: South Africa case study

by M. Mohutsiwa and C. Musingwini. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Formulae are developed that can be used for estimating the capital costs of developing underground bord and pillar, surface shovel and truck, and dragline coal operations. The error of magnitude level is -30% to +50%, which is suitable for a concept study.

These papers will be available on the SAIMM website

http://www.saimm.co.za

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http://dx.doi.org/10.17159/2411-9717/2015/v115n8a1

The physical ability of women in mining: can they show muscle?

by D. Botha*, and J.F. Cronjé*

Although women all over the world have been involved in mining activities for centuries, mining has always been considered a very masculine industry due to its heavily male-dominated workforce as well as the physicality of mining work. The mining industry has not been an obvious career choice and preferred place of employment for women; women were mainly employed in administrative and advisory positions.

Until 1994, women were legislatively prohibited from being employed in underground operations in South Africa, but the Mines Health and Safety Act, No. 29 of 1996, removed these restrictions. In addition, new mining legislation (the Mineral and Petroleum Resources Development Act, No. 28 of 2002) and the accompanying Mining Charter make specific provisions for the inclusion of women in core mining activities and require 10% of core positions to be filled by women.

This article voices perceptions of the physical ability of women employed in core mining positions. Findings are drawn from empirical work undertaken at platinum, phosphate, and copper mines. Quantitative and qualitative research paradigms are used. It is evident that women find it extremely difficult to perform mine work that requires physical strength and stamina. Practical recommendations, informed by the literature review and empirical findings, are made with the objective of contributing to the sustainable deployment of women in the mining industry.

Keywords core mining activities, mining industry, physical ability, sustainable deployment, women in mining.

Introduction For many years, mining has been considered the foundation of the South African economy.

Although the mining industry in South Africa is currently under considerable pressure and experiences various challenges, including escalating operational costs, electricity tariff increases, safety-related issues and associated production stoppages, poor productivity, labour unrest, and reduced demand both globally and domestically, it remains a key contributor to the national economy and development of the country (IDC, 2013, p. 9).

Until 1994, women were legislatively prohibited from meaningful participation in the mining industry and were denied access to skills and jobs, self-employment, and entrepreneurship. The newly elected democratic government in 1994 initiated substantial socio-political and economic transformation in South Africa. Nearly every sector in the country was transformed and re-shaped, The Journal of The Southern African Institute of Mining and Metallurgy

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* North-West University. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Mar. 2014 and revised paper received Nov. 2014.

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primarily through sector-specific legislation, but also through negotiations between government, industry, and labour unions to create and refine the so-called sector charters, which included the Mining Charter. The vision of the new regime, the African National Congress (ANC), was to ‘transfer power to the people and transform society into a non-racial, non-sexist, united, democratic one, and change the manner in which wealth is shared, in order to benefit all the people’ (ANC, 2007).

Transformation in the South African mining industry is governed by the provisions of the Mineral and Petroleum Resources Development Act, No. 28 of 2002 (RSA, 2002), which was promulgated and implemented on 1 May 2004, and the Broadbased Socio-economic Empowerment Charter for the South African Mining Industry (RSA, 2004) (hereinafter referred to as the Mining Charter), which was signed in October 2002 and formally published on 13 August 2004.

The broad objectives of the Act and the accompanying Mining Charter (and the amended Charter) are to rectify previous inequalities and disadvantages in the mining sector and to ensure equity, accessibility, and sustainability in the industry. The Act and the Charter make specific provisions for the inclusion of women in core mining activities, in terms of which the industry was supposed to reach a quota of 10% women in core mining activities by 2009. Core mining activities include, among other activities, mining, metallurgy, engineering, and geology (Harmony Gold Mining Company, 2008, p. 32).

In 2009, the Department of Mineral Resources conducted a thorough impact assessment to determine the progress made

The physical ability of women in mining: can they show muscle? against the Mining Charter objectives, as adopted in 2002, regarding transformation in the mining industry. This culminated in the Amendment of the Broad-based Socioeconomic Empowerment Charter for the South African Mining and Minerals Industry (RSA, 2010), which was launched in September 2010. The Amendment of the Charter set further requirements in terms of employment equity targets. Mining companies are required to reach a target of 40% historically disadvantaged South African (HDSA) representation in core and critical skills by 2015 and, in addition, a 40% HDSA representation in junior management levels by 2011, middle management levels by 2013, and senior/executive management levels by 2015 (Cliffe Dekker Hofmeyr and Reid, 2010). The term HDSAs refers to black people (African, so-called Coloured people and Indians), women, and people with disabilities (Nel et al., 2009, p. 79).

Although women all over the world have been involved in mining activities for centuries, mining is considered a very masculine industry due to its heavily male-dominated workforce as well as the physicality of mining work.

Furthermore, the nature of working in mines, and specifically underground, is hazardous and extensive training is required (Wynn, 2001, p. 34). The mining industry has therefore not been an obvious career choice or preferred place of employment for women; women are mainly employed in administrative and advisory positions.

Although well-intended, the introduction of women into the very male, ‘macho’ mining environment created and still creates challenges, not only for managers but also for male co-workers and the newly employed female mineworkers. On the one hand, female employees need a sound level of overall fitness to complete everyday work as well as to achieve independence and credibility in the eyes of their co-workers.

On the other hand, the mining industry is production-driven and employees are continually under pressure to reach production targets. Due to the physical differences between women and men, women often find it difficult to perform certain work activities and tasks and thus have a direct negative impact on production targets.

The purpose of this article is to voice perceptions of the physical ability of women employed in core mining positions.

Findings are drawn from empirical studies undertaken at platinum, phosphate, and copper mines.

Research objectives The specific objectives of this article are three-fold: ➤ To highlight issues regarding the physical ability of women employed in core mining positions according to the literature ➤ To determine issues regarding the physical ability of women employed in core mining positions according to the empirical findings ➤ To provide recommendations to address the issues identified so as to contribute to the sustainable deployment of women in the mining industry.

Research methodology Research approach A mixed-method research design was followed by applying both quantitative and qualitative research approaches.

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Empirical context Research participants The research setting was limited to the following three mines: an underground copper mine, an underground platinum mine, and an opencast phosphate mine. The mines were selected on an availability basis (convenient sampling).

For the purpose of quantitative research, the study population consisted of an availability sample of management as well as male and female employees working in core mining activities at the three mines. In total, 156 responses were received – 68 from the copper mine, 38 from the platinum mine, and 50 from the phosphate mine.

Purposive or judgemental sampling was used to select participants for the qualitative research. In total, 12 individual interviews and 19 group interviews (69 participants) were conducted. The researcher aimed to gain information from various operations; the participants were therefore selected from various categories of employment and mining disciplines.

Measuring instruments Quantitative data was collected by means of a structured questionnaire. Qualitative data was collected by means of individual interviews and group interviews. Both the individual interviews and the group interviews were semistructured, as an interview guide was utilized. Data collected was audio- and video-recorded and written notes were taken.

Research process The researcher formally requested permission from mine management to conduct research at the three mining companies. After permission was granted, a formal appointment was scheduled with mine management to explain the nature and extent of the research. In each research setting (mine), a contact person (the human resource officer targeted with managing women in mining issues) was allocated to the researcher to provide the necessary assistance and support during the research, which included distributing and collecting of the quantitative questionnaires, selecting appropriate participants for the individual and the group interviews, scheduling interviews, and organizing the underground field trip as well as visits to surface mining operations. Most of the individual and group interviews were scheduled between shifts in order not to interfere with the work responsibilities of the participants.

Ethical considerations, such as voluntary participation, informed consent, privacy, anonymity, and confidentiality, as recommended by Babbie and Mouton (2011, p. 520), were taken into account while conducting the research. The research was also approved by the Committee of Advanced Degrees within the Focus Area Social Transformation of North-West University.

Data analysis Quantitative data obtained through the questionnaires was analysed with the support and assistance of the Statistic Consultation Service of North-West University. The statistical software program SPSS 21.0 for Windows™ was used to analyse the data. Qualitative data obtained through the individual and group interviews and observations was analysed by means of conceptual (thematic) analysis.

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The physical ability of women in mining: can they show muscle?

Firstly, descriptive statistics and frequencies are presented, differentially in terms of the three mines included in the study. Descriptive statistics are reported per statement as mean values. The means can be interpreted as follows: ➤ Ratings of 2 and below indicate that the majority of the research participants disagreed or strongly disagreed with the indicator statement ➤ Ratings above 2 indicate that the majority of the research participants agreed or strongly agreed with the indicator statement. (The maximum response for each statement is 4.)

Secondly, the findings of a factor analysis, conducted to explore the factorial structure of the section, are reported and discussed. Thirdly, effect sizes were measured. Because an availability sample was used, p-values were not relevant and differences between means were examined for practical significance with effect sizes. Lastly, the findings of the qualitative enquiry are reported.

Limitations A limitation to the study lies in the accessibility of the mining sector as a research setting. Past research has shown that it is sometimes extremely difficult to interview employees and representatives of mining companies and to get them to fill in questionnaires. It was not an easy task to gain access to the mining companies. Several visits and exchanges of correspondence took place before permission was granted for the research. In addition, the platinum mine experienced many difficulties as well as labour unrest during the period that the research was conducted. Because of this, several interviews with management were postponed and eventually cancelled. Despite numerous attempts by the researcher, no quantitative responses (questionnaires) were received from the management target group of the platinum mine.

Furthermore, not all participants targeted for the individual and group interviews turned up for the meetings. Some of the participants could not stay for the full duration of the interviews due to work responsibilities and emergencies.

Others were drained and tired after shift work and wanted to depart for home as soon as possible to get some rest and take care of their family responsibilities before the start of their next shift. The researcher made use of existing skills, knowledge, and networks to overcome some of these problems.

Literature review Work in the mining sector is associated with difficult working conditions, and mining, especially underground, is considered one of the most physically demanding occupations (Schutte, 2011, p. 11). The nature of working in a mine, and in particular underground, is hazardous and extensive training is required. Many jobs also require a high degree of physical strength and endurance. Mine work includes, among other tasks, the ability to carry heavy objects and to work outside, underground, and in confined spaces, often in hot conditions for extended periods of time (Wynn, 2001,, p. 34).

Some of the work tasks are difficult to perform due to the physical differences that exist between women and men (Wynn, 2001, p. 33). Due to women’s smaller physical work capacity and physical strength, they may experience undue The Journal of The Southern African Institute of Mining and Metallurgy

physiological strain when performing prolonged and strenuous physically demanding tasks (Ashworth et al., 2004, p. 34). Research conducted by Badenhorst (2012) revealed that 43% of new female recruits are not fit for physically demanding work in mining. The author attributes this mainly to the following: ➤ Genetic predisposition of women. Women have physiological disadvantages when performing physically demanding work ➤ Workplace design. The designs of equipment, machinery, and the workplace still cater for the male population ➤ Lifestyle. Women tend to have a less active lifestyle than their male counterparts.

A study conducted by Schutte et al. (2012, p. 3), which assessed workplace stress associated with routine mining activities, confirmed that women experience significantly more physiological strain than men when performing mining tasks.

Furthermore, much of the equipment used in South African mines is designed overseas for use by men, who tend to be significantly taller than the average South African woman (Schutte, cited in Campbell, 2007). The ergonomic features of such mining equipment do not provide for the physiological make-up of women. When mines started to recruit women, equipment was still only suitable for men, resulting in both women and men having to share such equipment. Body dimensions are an important concept that should be taken into consideration in the design of mining equipment as well as its efficient operation (Campbell, 2007).

Mine work also requires manual handling of equipment and tools. These are designed to accommodate the size and strength of men (Zungu, 2011). Although many tasks within the industry are being automated, the industry will essentially remain labour-orientated, as it will always include manual tasks (Wynn, 2001, p. 34). Such tasks can often cause back injuries and musculoskeletal disorders for workers (Badenhorst, 2009, p. 63). In general, the manual material-handling capabilities of women are substantially lower than those of men, attributed primarily to differences in muscle strength. On average, a woman’s lifting strength is 60 to 70% of that of a man (Ashworth et al., 2004, p. 34).

According to Zungu (2011), there is a need to increase awareness and knowledge of the ’safe limits’ for women for handling mining equipment and tools. Furthermore, women cannot use the same techniques as their male counterparts to lift and handle heavy objects and materials. Women’s size and body build should be taken into consideration when appointing women in core positions (Zungu, 2011).

Gender differences also exist regarding the aerobic capacity of men and women. Aerobic capacity refers to the maximal oxygen uptake that provides a quantitative measure of a person’s ability to sustain high-intensity physical work for longer than five minutes (Ashworth et al., 2004, p. 34).

The aerobic capacity of women is typically 15 to 30% below the values of their male counterparts. This means that women work closer to their aerobic capacity than men and are thus more likely to become fatigued. Fatigue is operationally defined as the ‘reduced muscular ability to continue an existing effort’ (Ashworth et al., 2004, p. 34).

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The physical ability of women in mining: can they show muscle? tivity in the workplace and increase the risk of accidents and injuries occurring (Schutte, 2010, p. 54).

People working in mines, and in particular underground, are also exposed to extreme heat. The underground environment is dark and damp, and is characterized by an increase in temperature with depth (Singer, 2002, p. 1). In South African mines, work environments with a wet-bulb temperature higher than 27.4°C are considered to be hot and necessitate the introduction of practices to safeguard miners (Schutte, 2009). According to Ashworth et al. (2004, p. 35), all employees who work under ‘conditions conducive to heat stroke should be screened for gross or permanent heat intolerance by means of the standard heat tolerance screening test procedure. Heat tolerance screening assesses whether an individual can withstand high temperatures while doing physically demanding work and is used to protect individuals against the negative consequences of heat exposure, such as heat stroke and heat-related diseases (Benya, 2009, p. 56).

In the South African mining industry, heat tolerance screening consists of bench-stepping for 30 minutes in a climatic chamber at an external work rate of approximately 80 W in an environment with a dry-bulb temperature of 29.5°C and a wet-bulb temperature of 28.9°C. If the person’s body temperature does not exceed a given value at the end of the test, the person is classified as potentially heat-tolerant, indicating that he or she is fit to undertake physically demanding work in a hot environment (with wet-bulb temperatures greater than 27.5°C) (Schutte, 2009, p. 3).

Female employees are not given any privileges and must pass the same medical and screening tests as male employees (Singer, 2002, p. 2).

High occupational heat loads can lead to the following problems, among others: impaired work capacity; errors of judgement, with obvious implications for safety; lethargy and fatigue; and complications such as heat stress, which can lead to heat stroke, which is often fatal (Schutte, 2009, p. 1).

The following personal risk factors, among others, may reduce an individual’s tolerance for heat stress: age, obesity, state of hydration, use of medication and drugs, gender, and acclimatization state (Zungu, 2011). Furthermore, gynaecological conditions and pregnancy can also affect the way in which women handle heat stress (Zungu, 2011).

According to Badenhorst (2009, p. 59), a female employee can do any job that she is qualified to do, provided that she meets the requirements inherent for the specific job.

Furthermore, an employee should not be employed in a job or conduct tasks for which he or she is not medically fit or if he or she does not have the required physical and functional capabilities. The health and safety of the employee and coworkers should not be compromised (Badenhorst, 2009, p.

59). Therefore, Badenhorst (Badenhorst, 2009, p. 70) suggests that a programme be established to ensure that minimum medical requirements are met by employees, which include the establishment of minimum standards for fitness, comprising the following steps:

Step 1: Occupational health risk assessment A clearly defined occupational health risk profile should be created for each occupation by identifying all relevant health hazards and the degree to which workers are exposed to these hazards.

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Step 2: Man-job specification The risks for each and every occupation should be documented, which should cover both the inherent requirements of the jobs and the expected hazard exposure(s).

Step 3: Setting standards for medical surveillance A medical practitioner should set medical standards for each of these occupations based on the risk profiles. These should include standards for physical and functional ability required to perform certain jobs safely. A test battery to conduct and measure these abilities should be established.

From the above discussion, it is evident that various factors need to be considered when integrating women into the core business of mining in order not to compromise the health and safety of both female employees and their male colleagues.

Empirical findings and discussion Physical ability of women employed in core mining activities Statements were included to determine the perceptions of the three target groups, namely men and women employed in core mining positions and management, of the physical ability and capability of women employed in core mining positions. The descriptive statistics are presented in Table I for the three target groups at each of the three mines. To avoid duplication, the results are explained according to factor analysis. Factor analysis can be defined as a technique for identifying groups or clusters of variables (statements) (Field, 2005, p. 619).

Three factors (groups of variables) were extracted using Kaiser’s criteria (Field, 2005, p. 652) that explain 39.92% of the total variance in the section on Physical ability. All statements have satisfactory factor loadings of above 0.3.

Factor loadings represent the intercorrelation between the statements (variables). According to Field (2009, p. 644), an absolute value of 0.3 is regarded as important. Questions 1, 2, and 5 loaded on Factor 1 (Capability), questions 3 and 4 loaded on Factor 2 (Effectively), and Question 6 loaded on Factor 3 (Differential). It is evident from the quantitative results that the majority of the participants across all three mines agreed with the statements contained in the Capability (mean = 2.74), Effectively (mean = 2.77), and Differential (mean = 2.19) factors. On average, the perceptions with respect to Factor 1: Capability are that (i) women are physically less capable than men, (ii) some mining tasks can only be done by men, and (iii) temperatures in the workplace are regarded as a major problem for women. With respect to Factor 2: Effectively, (i) women have the physical ability to perform their daily tasks effectively, and (ii) they do not have a problem with working in confined spaces. With respect to Factor 3: Differential, it is perceived that (i) women should be treated differently to their male co-workers in the workplace.

The Capability factor shows a Cronbach’s alpha coefficient of 0.68, which could be regarded as an acceptable reliability.

Cronbach’s alpha is one method of estimating the reliability and internal consistency among the statements (Field, 2009, p. 675). According to Field (op. cit., p. 668), the Cronbach’s alpha value could realistically be below 0.7. The Effectively factor shows a Cronbach’s alpha coefficient of 0.54, which The Journal of The Southern African Institute of Mining and Metallurgy

The physical ability of women in mining: can they show muscle?

Table I

Participants’ perceptions regarding the physical ability of women working in core mining activities Copper mine Indicator statement

Males in core

Females in core

Phosphate mine Management

Males in core

Females in core

Platinum mine Management

1.

Women are physically less capable 3.50 2.56 3.19 3.13 2.26 2.92 than men 2.

Some mining tasks can be done 3.56 2.62 3.31 3.31 2.84 3.08 only by men 3.

I (women) have the physical ability to 2.63 3.41 2.75 2.59 3.30 2.67 perform my (their) daily tasks effectively 4.

I (women) find it easy to work in 2.31 2.53 2.00 2.00 2.39 2.50 confined spaces 5.

Temperatures in the workplace are 2.69 2.30 2.44 3.00 2.00 1.90 regarded as a major problem for women 6.

Women should be treated differently 2.00 2.00 1.75 2.00 2.71 2.25 than their male co-workers in the workplace Mean scores of above 2 indicate that the majority of the research participants agreed or strongly agreed with the indicator statement.

Males in core

Females in core

3.25

2.50

3.13

3.05

3.07

3.32

2.67

2.63

2.47

2.25

2.40

2.15

Table II

Comparison of the three target groups of the different mines regarding physical ability Mine

Mean

Women Standard deviation

Mean

Factor 1: Capability

Phosphate 3.18 0.64 2.39 Copper 3.25 0.55 2.49 Platinum 2.90 0.54 2.65 Factor 2: Phosphate 2.29 0.79 2.83 Effectively Copper 2.47 0.64 2.99 Platinum 2.81 0.57 3.02 Factor 3: Phosphate 2.00 1.12 2.71 Differential Copper 2.00 0.97 2.00 Platinum 2.38 1.20 2.16 (a) Small effect: d = 0.2, (b) medium effect: d = 0.5, and (c) large effect: d = 0.8

could be regarded as having relatively low reliability. Factor 3 consists of one item only, therefore Cronbach’s alpha is not applicable.

From Table II, it follows that the effect sizes for the Capability factor show a medium and large effect, indicating that, on average, the participants from the male and management target groups are more in agreement with the indicator statements contained in the Capability factor than the female target group themselves. The effect sizes for the Effectively factor indicate that the participants from the female target groups of the phosphate and copper mines are more in agreement with the indicator statements contained in the factor than the participants from the male and management target groups at these mines. A medium effect is evident from the effect sizes of the phosphate mine for the Differential factor, indicating that, on average, the female participants from the phosphate mine feel that they should be treated differently from their male co-workers in the workplace, while this view is not supported by the research participants in the male and management target groups.

The Journal of The Southern African Institute of Mining and Metallurgy

Management

Effect sizes

Standard deviation

Mean deviation

Standard

Women vs Men

Women vs Management

0.70 0.66 0.53 0.61 0.63 0.58 0.96 0.80 0.90

2.68 2.98

0.37 0.74

0.41 0.67

2.58 2.38

0.42 0.53

2.25 1.75

0.62 0.68

1.12 1.16 0.47 -0.67 -0,80 -0.36 -0.64 0.00 0.18

-0.39 -0.96 -0.49 -0.31

Perceptions of women’s performance and position in core mining activities Questions were also included in the questionnaire to determine whether women feel confident in performing core mining activities and to verify perceptions of male co-workers and management regarding women’s confidence in performing core mining activities.

Figure 1 shows that the majority of the women in all three mines (copper mine: 67.9%; phosphate mine: 57.9%; platinum mine: 68.4%) feel confident in performing their respective work activities. Only a few participants (copper mine: 3.6%; phosphate mine: 15.8%; platinum mine: 15.8%) indicated that they do not feel confident at all.

The male and management participants were asked to give their opinions on whether women feel confident in performing the following core activities: driving a locomotive, driving a winding engine, operating a conveyer belt, using heavy and/or vibrating power tools, driving a winch, and operating a shift. The results are presented in Table III. VOLUME 115

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Men Factor

The physical ability of women in mining: can they show muscle?

Figure 1 – Female participants’ perceptions regarding their confidence to perform work activities

The results reported in Table III show that the majority of the participants in the male and management target groups thought that women are confident in performing all mentioned activities, with the exception of the use of heavy and/or vibrating power tools. This view is supported by the findings from the literature review as well as the qualitative enquiry.

Perceptions and major concerns regarding the physical ability and capability of women employed in core mining positions Equipment, tools, and work units The data obtained from the qualitative enquiry revealed that women are employed in all sections at the mines, underground as well as on the surface, and fill positions such as engineer, geologist, electrician, artisan, fitter, boilermaker, and operator of heavy machinery. They are also involved in technical and mechanical mining operations. However, pregnant women are not allowed to enter work areas in which they could be exposed to radiation. The participants also indicated that no mining equipment and tools are banned from use by women, with the exception of load haul dump (LHD) machines. When female employees initially operated these machines, many complaints were lodged due to the vibrations caused. The female participants indicated that vibration caused by the LHD machines had a negative impact on their health and interfered with their menstrual cycles. As a result, women are no longer allowed to operate LHD equipment.

Although female employees are employed as rock drill and winch operators, the female participants reported that these types of equipment are too difficult for women to operate because they are too heavy. Furthermore, the participants from the management target group at the phosphate mine indicated that they have experienced problems when women operated rubber dozers, as noted in the following comment: What happens, because that dozer drills the whole time, it interferes with females’ monthly period ...There are complaints, they go to medical doctors, the doctors have picked it up, it seriously disturbed them, it makes them seriously sick, you can check at a couple of mines we have done our homework, they experience the same problem. They rather put them in the track dozer and not in the rubber dozer, because the rubber dozer is the one that vibrates the whole time. I have had four cases of ladies who had medical problems with regard to that and we had to move them around to the track dozers instead of the rubber dozers.

Perceptions and major concerns of management From the interviews and focus group discussions held with the participants in management positions, it became clear that although job reservation does not exist and women are employed in all sections at the mines, women do not have the physical strength and ability to be employed in all core mining positions. According to the participants, it is important to consider job specifications and requirements when appointing women in positions that require physical strength. Women are often appointed in core mining positions without having the physical ability to cope with the requirements of these positions; male employees have to assist and support their female colleagues to do their jobs properly. This often leads to frustration on the side of male co-workers and contributes to a negative attitude towards women in mining. Male co-workers are often unwilling to assist their female colleagues because they feel that since women appointed in these positions receive equal salaries, they should do their jobs on their own. The following quotations provide an indication of some participants’ opinions with regard to the physical ability of women employed in core mining positions: You know I think it’s about the company’s recruitment and selection procedures and plans. Every human being has his own strengths and weaknesses. There are women

Table III

Male and management participants’ perceptions regarding women’s confidence in performing core mining activities Indicator statement Do you think women are confident in performing the following activities?

Male in core Copper mine

Phosphate mine

Management Platinum mine

Copper mine

1.

Driving a locomotive 3.00 3.12 3.43 3.07 2.

Driving a winding engine 3.00 2.54 2.80 2.93 3.

Operating a conveyer belt 3.06 3.47 3.21 2.94 4.

Using heavy and/or vibrating power tools 1.69 2.23 1.60 1.94 5.

Driving a winch 2.75 3.00 1.87 2.81 6.

Operating a shift 3.06 3.36 3.19 3.00 Mean scores of above 2 indicate that the majority of the research participants agreed or strongly agreed with the indicator statement.

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Phosphate mine

3.08 3.22 3.08 2.45 3.00 3.00

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The physical ability of women in mining: can they show muscle?

Perceptions and major concerns of male co-workers Data obtained from male employees working in core mining positions revealed frustrations regarding the physical ability of women, production targets, crisis situations, attitudes of women, and confrontations with husbands/boyfriends.

The male participants clearly indicated that women often lack the physical strength to perform certain work activities.

It was noted that women are able to perform light duties, but experience difficulties when operating heavy machinery and lifting heavy objects. Furthermore, it was indicated that women often do not have the stamina to perform certain work activities. According to the male participants, women’s bodies are not made to perform hard physical mine work.

Male co-workers often have to support and assist female employees when they do not have the physical strength to perform their work. Some male participants said that they are willing to assist women who are willing to work, but they are negative towards women who do not show a willingness to do the work for which they were appointed. Others get agitated when they have to assist women or do/complete their work for them. The following comments illustrate these points: Some work women can do, and some not. Some heavy objects they cannot lift. They find it difficult to work in confined spaces. It is not all the women who can use the safety harness to work at heights. Some of the duties at my section, I just do it by myself, because I see that they find it tough.

When [it comes] to physical ability, I don’t think they have that power to work as a man, because most of the [work is] underground. You need physical strength, so if you don’t have that strength you won’t make it to do that job correctly, because some of them you can see that she is trying, but she doesn’t have the power to do that.

From my experience: When the work gets tough, the woman move to one side and let the men do the work.

Men do get agitated when they have to assist women.

We have to speak to them and convince them that these are women.

The Journal of The Southern African Institute of Mining and Metallurgy

The mining industry is production-driven and highly focused on reaching production targets. According to the male participants, the inclusion of women in a mining team has a severe impact on productivity, as they feel that women are not physically able and capable of pulling their weight in the team and they often lack physical strength and stamina.

The male participants reported that they prefer to work with an all-male team due to the fact that mine work should always be done against time in order to reach production targets. If certain tasks are not performed well, delays result.

Furthermore, pregnancy can affect the productivity of the team. When pregnant, women are not allowed to perform certain work activities; they are employed elsewhere in the company. The company does not always appoint another team member in the place of the pregnant woman, and the team thus has to rely on fewer workers to complete the same tasks. This also has a severe impact on reaching production targets. These points are confirmed by the following responses: Last week I was talking to this other supervisor in the plant. He told me that in his shift, he prefers to have a male than to have a female, because he knows when he has 10 males he can do his job quicker. If he has nine men and only one woman he knows there will be some delays, because there are some jobs that a woman cannot do.

You know sometimes the reason why men help women to do their job, it’s because of time or because of the work we are doing. Maybe it’s too much for us and we must knock off, maybe at 14:00. So women are not so much strong, so we must make it snappy. So they must sit down there so that we can do it.

Due to the fact that mine work is associated with a high risk of accidents, male employees often prefer to work with men because they feel that men can act faster in crisis situations, as was pointed out by one of the participants: Women are not faster than men. For example, when there is a fire on top of the plant, something is burning, the woman cannot pick up the fire extinguisher and rush to the fire, but the man can easily do that.

The participants from the platinum mine indicated that male co-workers often experience problems with women’s attitudes. Although they acknowledge that some women are willing and capable to do the work for which they are appointed, others want to be ‘treated like ladies’ in the workplace. The following quotations provide an indication of the participants’ opinions in this regard: Women are always complaining. They want us to treat them special, but they don’t want to do the work. They complain that the work it is too heavy. And underground, there is no easy job there. Everything is very hard.

Some women they want to be treated like ladies; […] we don’t have such a kind of chance to treat them like ladies, because we are always in a hurry. Underground everything is done against time and against production – the mining company wants production the whole time.

The ladies they want to be treated like glass, eggs.

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that are stronger than men and there are men stronger than women. It is important when a position/job becomes available to look at the requirements of this position/job and then evaluate how you will determine whether candidates are physically strong enough to fill this position. In the past, we have often made mistakes with the appointment of women by not looking at the requirements of the position/job. We appointed thin, fragile women who could not cope with the requirements of the position/job. The male colleagues then often have to assist women who are not physically strong enough to do their jobs and they then become easily frustrated and annoyed with their female colleague.

It depends on the type of work they do. For example, women find it difficult to do mechanical type of work that requires physical strength. On the other hand, women perform well if they do work that requires little or no physical hard labour. But as soon as they start to do heavy work or work that need some physical strength or power, they cannot do it.

The physical ability of women in mining: can they show muscle?

There was this other guy at the station. He came to me straight and asked me why do I give his girlfriend a hard job. So do you see what is this thing causing? Men are complaining to us or maybe they want to fight us.

Perceptions and major concerns of women employed in core mining positions Although the female participants indicated that they have the physical ability to do their jobs well, they admitted that the work is tough and not easy to perform, especially underground. It was also reported that, on average, the women are willing and able to perform their jobs well; however, the women admitted that they do not always have the physical strength, power and stamina required for specific positions. They reported that some male co-workers are willing to assist them, while others are unwilling and would rather watch them suffer than be of assistance. Women want to prove themselves and often neglect their bodies to do their jobs well. The following quotes provide indications of the female participants’ opinions regarding the physical ability of women employed in core mining positions and the constraints experienced: I don’t have the steam to work at the position that I am working at. I am not strong enough. The job of mine is too hard. We are sweating underground.

The loco is like a train, nè? It’s hard to operate. The steering wheel and everything is hard. The brakes. And to be on it every day, yô, it is hard. When you go on period you have some pains. Your back it pains. And that thing, it vibrates. I’m on it eight hours every day.

Yes, I have the physical ability to do my work on my own. It’s just, I have to make a plan to go out on my own. I find that some of the male workers are using power to do the job and I find that I have to come with a plan to make it simple.

The empirical results confirm the findings of the literature review, which suggest that female mineworkers are at a disadvantage in terms of ability to perform mine work that requires physical strength and stamina. They experience physiological strain when performing prolonged and strenuous physically demanding tasks.

Conclusion and recommendations It is evident from the literature review, and confirmed by the empirical findings, that various factors (such as physical strength and fitness, heat tolerance, body dimensions, and ergonomic features of mining equipment) need to be considered when appointing women in core mining positions.

Although physical fitness tests and heat tolerance screening are carried out prior to the appointment of women in core mining positions, and regardless of the kind of mining (underground or opencast), the empirical findings confirm that women are still appointed in positions entailing work that they find extremely difficult to perform, such as the operating of heavy machinery (LHD equipment, rubber dozers, rock drills, and winches), or performing mine work that requires physical strength and stamina and using heavy and/or vibrating power tools. Furthermore, due to women’s inability to perform mine work that requires physical strength and stamina, management and male co-workers experience unique frustrations and challenges.

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In light of the above, the following recommendations, informed by the literature review and the empirical findings, are made to contribute to the sustainable deployment of women in the mining sector: ➤ There is an urgent need to conduct research on and identify specific issues that could cause physiological strain to the female mineworker in order not to compromise their health and safety. Creating awareness of these issues is of utmost importance ➤ Mining companies should review their recruitment and selection procedures to maximize the fit between female mineworkers and their specific positions in the mines ➤ An employee should not be appointed in a position or conduct tasks for which he or she is not medically fit or does not have the physical and functional capabilities ➤ An employee should be appointed in a position only if he or she meets the requirements for that specific job ➤ Regular training courses should be conducted to educate women on correct biomechanics when operating heavy machinery and using heavy and/or vibrating power tools ➤ Fitness and conditioning programmes should be developed and established to assist women in handling their everyday work activities ➤ Regular diversity training and workshops focusing on aspects regarding the female miner, such as the physical strength and stamina of women, and physiological aspects related to the female body, such as menstruation, pregnancy, and birth, should be presented to male and female employees to foster a work environment in which differences (in terms of gender) are respected.

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HOFMEYR, C.D and REID, A. 2010. The Revised Mining Charter: How does it affect you? A comparative analysis. www.polity.org.za/.../the-revisedmining-charter-how-does-it-affect-you-a-comparative-analysis-2010-0920 [Accessed 26 November. 2010].

advanced metals initiative

Nuclear Materials Development Network Conference 28–30 October 2015 Nelson Mandela Metropolitan University North Campus Conference Centre, Port Elizabeth

A n n o u n c e m e n t

academia, the nuclear industry and government in order to share and debate the latest trends, research and innovations, specifically in the field of nuclear materials, local mineral beneficiation and advanced manufacturing related to these materials.

Keynote speakers to be invited include international specialists in the field of nuclear materials and mineral beneficiation.

The event will kick off with a Student Seminar on 28 October 2015 covering the full scope of the four AMI Networks.

A visit to the world class High Resolution Transmission Electron Microscopy Laboratory and Mechanical Engineering Department of NMMU (eNtsa) will form part of the conference.

An opportunity that cannot be missed!

BACKGROUND Through its Advanced Metals Initiative (AMI) the South African Department of Science and Technology (DST) promotes research, development and innovation across the entire value chain of the advanced metals field.

The goal of this initiative is to achieve sustainable local mineral beneficiation and to increase the downstream valueaddition of advanced metals in a sustainable manner. The achievement of this is envisioned to be through human capital development on post-graduate and post-doctoral level, technology transfer, localization and ultimately, commercialisation.

The AMI is comprised of four networks, each focussing on a different group of metals. These are the Light Metals, Ferrous and Base Metals, Precious Metals and Nuclear Materials (zirconium, hafnium, thorium, tantalum, niobium and nuclear ceramic materials such as SiC, B4C, ZrN, ZrB2, ZrC, Si3N4, etc.).

The AMI Nuclear Materials 2015 Conference aims to bring together stakeholders from the mineral sector,

our future through science

WHO SHOULD ATTEND Stakeholders from private industry, science councils, academic institutions as well as the nuclear community at large are encouraged to attend.

EXHIBITION/SPONSORSHIP Sponsorship opportunities are available. Companies wishing to sponsor or exhibit should contact the Conference Co-ordinator.

For further information contact: Head of Conferencing, Raymond van der Berg SAIMM, P O Box 61127, Marshalltown 2107 Tel: +27 11 834-1273/7 ·Fax: +27 11 833-8156 or +27 11 838-5923 E-mail: raymond@saimm.co.za · Website: http://www.saimm.co.za

http://dx.doi.org/10.17159/2411-9717/2015/v115n8a2

Transformation in the South African mining industry – looking beyond the employment equity scorecard

by N.V. Moraka* and M. Jansen van Rensburg*

Going beyond transformation claims contained in employment equity scorecards and industry compliance reports, this article provides qualitative insight into the initiatives employed and challenges experienced by mining companies in a quest to transform the mining industry. Perceptions expressed during in-depth interviews with 10 senior executives showed that the assumption that mining companies are reluctant to transform is erroneous. Results from this study suggest buy-in and commitment to transformation. This article describes specific initiatives undertaken by mining companies to transform. The most notable initiatives include staff recruitment efforts to appoint historically disadvantaged South Africans (HDSAs), staff development initiatives, as well as community development. The findings furthermore contextualize the challenges experienced by industry participants in their quest to transform. Despite accusations that industry participants are not taking responsibility for the implementation of the transformation agenda, government needs to recognize that they too have a role to play and need to appreciate and assist in the current challenges experienced by the industry.

Keywords transformation, mining charter, historically disadvantaged south Africans, black-economic empowerment.

Introduction ‘We need a mining sector that works. Mining employs over half a million people. It is the biggest earner of foreign exchange in our country. It also contributes about 20 billion rand directly to tax revenue. Mining also makes a far larger contribution as a buyer of goods and services, and a supplier of inputs to other sectors of our economy and other economies around the globe’ said President Jacob Zuma during the 2014 State of the Nation Address (Zuma, 2014). However, despite apparent commitment from government, enforced compliance with social and labour plans, regulations and Mining Charter targets, the South African mining industry is slow to gain local and international investor trust (Deloitte & Touche, 2013; Mashego, 2013).

Industry players acknowledge that their context has changed – creating various challenges (Davis, 2014). In the past few years, such challenges included subdued commodity prices, increased working costs, constrained infrastructure, and high labour The Journal of The Southern African Institute of Mining and Metallurgy

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costs, coupled with poor levels of productivity, strained labour-management relations, an uncertain regulatory environment, and the inevitable maturation of the industry (Deloitte & Touche, 2013; Davenport, 2014). Miningaffected communities also became more cognisant of their rights, and politicians more vocal in their expectations from their relationships with the mining giants (Davis, 2014). One of the main issues that contributed to the difficult operating conditions was claims from organized labour and Minister Shabangu that this industry is too slow to transform.

Specific claims were made in terms of employment equity, fair salaries and wages for mineworkers, housing and living conditions, health and safety issues, and general working conditions (Limpitlaw et al., 2005; Miningmx, 2013; Shabangu, 2010). These claims were followed by what Gill Marcus, Reserve Bank Governor, explained as ‘unprotected strike action [that] has escalated into an uncontrolled, violent and unlawful landscape, led by a mob mentality in the absence of formal and recognised leaders’ (Mavuso, 2013).

In light of the stern warnings from government that 2014 is the deadline for mining companies to improve housing and living conditions of mineworkers and to achieve a number of targets (Miningmx, 2013; Zuma, 2014), the purpose of this article is to provide insight into the status of transformation in the mining industry. Specific attention will be given to current initiatives undertaken to drive transformation, challenges experienced, and finally to identify the barriers to transformation. The scope of this article goes beyond transformation claims contained in employment equity scorecards or the findings presented in industry compliance

* University of South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Oct. 2014 and revised paper received Jan. 2015.

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Synopsis

Transformation in the South African mining industry reports (see, for example, Mitchell, 2013 for a review).

Instead the article will report current initiatives and qualitative perceptions about the challenges experienced in the quest to transform the industry. The literature review will focus on relevant transformation regulation and legislation to set the legislative context. Progress on transformation in the mining industry reported elsewhere will also be considered.

The literature review is followed by a description of the methodology used to explore the status of transformation and then the results and findings are presented. Finally, conclusions are drawn, which aim to set recommendations for industry stakeholders.

Transforming the mining Industry The Mining Charter is the policy instrument to effect transformation of the South African mining sector (Department of

Mineral Resources, 2009). Despite being considered a priority by government and industry, the evidence on the extent of transformation in the mining industry is, at best, unreliable (Mitchell, 2013). This is curious, as industry stakeholders, including the Chamber of Mines, the South African Mining Development Association, and the National Union of Mineworkers negotiated and signed the Mining Charter. This charter is furthermore complemented by relevant legislation for mining of mineral resources, listed in Table I. Table I supports the notion that transformation is guided by clear legislation, transformation targets, and objectives.

The shared purpose of these Acts is to create true democracy and a non-racial country (Esterhuyse and Nell, 1990). In support of the constitution, the legislation and Acts furthermore strive to recognize human rights, improve the quality of life, and promote equality for historically

Table I

Chronological list of legislation affecting mining and BEE in SA Act number

Act name

Brief description

Date of commencement

Act 16 of 1967

Mining Titles Registration Act

1 October 1967

Act 20 of 1967 Act 61 of 1973 Act 78 of 1973

Mining Rights Act Companies Act Occupational Diseases in Mines and Works Act Close Corporations Act

Regulates the registration of mining titles and other rights connected with prospecting and mining.

Deals with the issue of unwrought precious metals.

Governs the formation and regulation of companies in SA. Deals with the compensation for diseases contracted by persons employed in mines and works.

Governs the formation and regulation of Close Corporations . in SA Created the Diamond Board and instituted regulations for the trade in diamonds.

Provides for the continuation of MINTEK (Council of Mineral Technology) and its management by a board.

Regulated the mining industry until May 2004.

Act 69 of 1984 (GG 9285, 1; 4 July 1984) Act 56 of 1986 (GG 10291, 1; 25 January 1986) Act 30 of 1989 (GG 11783, 1; 31 March 1989) Act 50 of 1991 (GG 13253, 1; 22 May 1991) Act 66 of 1995 (GG 16861, 1; 13 December 1995)

Act 29 of 1996 (GG 17242, 1; 14 June 1996) Act 108 of 1996 (GG 17678, 1; 18 December 1996) Act 55 of 1998 (GG 19370, 1; 19 October 1998) Act 4 of 2000 (GG 20876, 1; 9 February 2000) Act 28 of 2002 (GG 23922, (1; 10 October 2002)

Diamonds Act Mineral Technology Act Minerals Act Labour Relations Act

Mine Health and Safety Act Constitution Act Employment Equity Act Promotion of Equality and Prevention of Unfair Discrimination Act Mineral and Petroleum Resources Development Act

Amended SA’s labour legislation to take account of the provisions of the final constitution, regulate trade unions and streamline procedures for resolution of employment disputes.

It also created the Labour Court and the Labour Appeal Court.

Deals with the protection of the health and safety of employees in a mining operation.

Contains a Bill of Rights that protects all South Africans.

Achieves equity in the workplace by eliminating unfair discrimination. Also adopts affirmative action measures.

Gives effect to the Equality clause (S 9) of the Constitution.

Transferred mineral rights from private holders to government as guardian of peoples of SA and makes special provision to benefit historically disadvantaged persons.

Act 53 of 2003 (GG 25899, Broad–Based Black Established a legislative framework for the promotion of 1; 9 January 2004) Economic Empowerment Act Black Economic Empowerment.

Act 49 of 2008 (GG 36523, Mineral and Petroleum Provides a framework for the regulation of associated minerals 1 31 May 2013) Resources Development including guidelines for the partitioning of rights. This Bill Amendment Act furthermore enhances provision relating to the regulation of the mining industry through beneficiation of minerals, the promotion of national energy security in order to streamline administrative processes and to align it with previous Acts.

Act 46 of 2013 (GG 37271, Broad-Based Black Economic Inserts new definitions and amends others, provides for 2; 27 January 2014) Empowerment Amended Act remuneration of council members, promotes compliance, incentive schemes to support black-owned enterprises, establishes BEE commission.

Source: Department of Mineral Resources, (2014); Rungan et al. (2005)

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1 October 1967 1 January 1974 1 October 1973 1 January 1985 1 October 1986 1 August 1989 1 January 1992 11 November 1996

14 June 1996 4 February 1997 1 December 1998 16 June 2003

1 May 2004

21 April 2004 01 January 2013

27 January 2014

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growth, and development of the mining industry (Shabangu, 2010). The amended 2010 Mining Scorecard served to supplement the amended Mining Charter and contained more measurable items (Miningmx, 2011). The revised charter sets a target of 26% black ownership of South Africa’s mining assets by 2014, as before, and adds that HDSAs should constitute 40% of the total at all levels of management of mining companies. At present, the DMR does not issue neworder mining rights or grant mining licences unless companies are BEE-compliant and have the necessary BEE credits and allotments in place (Miningmx, 2013). In 2013, a redrafted black economic empowerment bill was passed in the National Assembly. This bill will eventually stipulate entirely new BEE codes which are likely to be vastly different from the milestones in the Mining Charter.

Transformation progress In delivering her budget speech in May 2013, Minister Shabangu emphasized that mining houses will implement government’s transformation agenda – ‘come hell or high water’. Minister Shabangu directly confronted mining houses about the slow progress made on the BEE front. ‘[Every] other stakeholder suffered from a case of parochial amnesia in terms of their responsibility for the implementation of this transformation agenda’, Shabangu said. ‘We ended up with widely varied accounts on the extent or otherwise of the progress that has been made in this regard’ (Miningmx, 2013). However, mining analysts, lawyers, and industry players have different views about the industry’s level of compliance, particularly in respect of BEE codes (Miningmx, 2013). In general, there seems to be consensus that most South African mining houses have largely met the transformation objectives and that they will meet the stipulated 26% black ownership target. The way in which government will measure compliance with that target is, however, uncertain (Davenport, 2014). A frequently cited problem is the current disjoint between different definitions in relevant Acts, frameworks, and scorecards. At a fundamental level, the definition of historically disadvantaged individuals contained in the MPRDA Act/Mining Charter, for example, includes white women, but the BEE codes regard only black, coloured, and Indian people as previously disadvantaged (DMR, 2009; Miningmx, 2013). These definitions directly impact the calculation of BEE targets and such ubiquities create widespread dismay (Miningmx, 2013; Rungan et al., 2005; Tupy, 2002).

Despite the confidence that ownership targets will be reached, there is less certainty regarding the 40% employment equity target at all levels, especially management. Although mining houses claim that they will exceed targets at the lower management levels, they blame skills shortages in management, critical, and core skills as the main reason for not meeting targets at more senior levels (Delloite & Touche, 2013; DMR, 2010; Healing, 2012; Rungan et al. 2005). According to the Landelahni mining report, the mining industry is competing for scarce skills with infrastructure, manufacturing, and other local industries as well as the global mining industry. Reasons cited for the skills shortage are declining numbers of graduates in miningrelated qualifications, high HDSA staff turnover, and retirement (Healing, 2012; Landelahni, 2013). In 2008, the VOLUME 115

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disadvantaged South Africans (HDSAs) (Booyens, 2006; Swart, 2003). The Employment Equity Act, for example, enacted affirmative action measures with the aim of obtaining equity in the workplace (Republic of South Africa, 1998; Thomas, 2002). The enactment of the Skills Development Act aims to ensure the development of critical, core management skills of HDSAs in the workplace. The Broad-Based Black Economic Empowerment (BBBEE) Act followed as an intervention to address the apartheid legislation that prevented HDSAs from fully partaking in the economy (Burger and Jafta, 2006). This Act seeks to obtain a balanced strategy that addresses ownership, management, employment equity, skills development, preferential procurement, and enterprise development.

Specific to the mining industry, the Mineral and Petroleum Resources Development Act (MPRDA) documents and enables the transformation of national mineral and mining policies (Republic of South Africa, 2002; Van der Zwan and Nel, 2010). The MPRDA furthermore establishes that the mining sector has a duty to guarantee that exploitation of minerals shall benefit the economy, adhere to corporate social responsibility issues, safety, health, skills development, and provision of employment opportunities for HDSAs (Chamber of Mines, 2007). In support of this Act, mining companies were advised to introduce social plans, which are aimed to benefit the wider economy. Social plans were to be accompanied by an exploration plan (work plan), financial plan, mining plan, labour plan, environmental plan, empowerment plan, as well as a marketing plan (Cawood, 2004, p. 58).

Transformation in the mining industry is furthermore informed by the Broad-Based Socio-Economic Empowerment Charter for the South African mining and minerals industry (DMR, 2010; Fauconnier and Mathur-Helm, 2008). This charter addresses ownership and employee transformational targets and aims to promote equitable access to the nation’s mineral resources for all the people of South Africa; substantially and meaningfully expand opportunities for HDSAs to enter the mining and mineral industry and benefit from the exploitation of the nation’s mineral resources; utilize the existing skills base for the empowerment of HDSAs; expand the skills base of HDSAs in order to serve the community; promote employment; and advance the social and economic welfare of mining communities (DME, 2004). Initial transformation targets were captured in a Mining Scorecard that was officially launched in 2003 (Cawood, 2004; Fauconnier and Mathur-Helm, 2008). The 2003 Mining Scorecard was informed by the elements of the 2002 Mining Charter and also developed to be in line with the BBBEE Codes of Good Practice. While the Mining Charter expanded on ‘how to do it’, the Mining Scorecard illustrated ‘how companies will be evaluated’ (Rungan, Cawood, and Minnitt, 2005, p. 740).

This scorecard was, however, criticized because it was merely a checklist with ‘yes or no’ options, and concerns were raised over the practicability of measurement scales (Rungan et al., 2005). Additionally, a compliance assessment done during 2009 showed that the industry was not fully compliant and some targets were not met. As a result, the Mining Charter was amended in 2010 to provide more measurable items, scales, and targets. Indeed, it was the vision of the 2010 Mining Charter to facilitate sustainable transformation,

Transformation in the South African mining industry Landehlani mining survey revealed that local mining graduates’ pass rate was 13% compared to the expected 25% throughput rate for four-year programmes. South Africa is also experiencing a high shortage of well-qualified, competent, and experienced artisans and professionals in the mining sector (Landelahni, 2008). This explains the provision of attractive bonuses, spiralling salaries, and retention packages by mining companies to retain HDSAs possessing these attributes (Engdahl and Hauki, 2001).

Industry players furthermore argue that it is challenging to address the skills shortage given the deficiencies of the Mining Charter and unrealistic targets set (Davenport, 2014; Miningmx, 2013; Mokoena, 2006; Schoeman, 2010; Tupy, 2002). The skills shortage will thus remain an issue as long as there is ineffective leadership for driving transformation, inability by mining companies to identify and manage a talent pool, and the broad transformation legislation (Esterhuyse, 2003).

Notwithstanding these challenges, industry players are aware that their context has changed and claim to embrace the concept of transformation. They moreover publicly declare that they are investing the required resources to not only comply with legislation, but rather to achieve true change (Davis, 2014). This article will consider current initiatives undertaken to drive transformation and challenges experienced in the quest to identify the barriers to transformation. The research questions of the study were the following: ➤ What is the progress made in transforming the mining industry, and what initiatives have been put in place by mining companies? ➤ What are the challenges and barriers to transformation in the South African mining industry?

The methodology used to explore the status of transformation is reviewed below.

analysis was used to analyse transcribed verbatim data. A summative approach to qualitative content analysis was used, which entailed counting and comparing quotations, keywords, or paragraphs followed by the interpretation of the underlying context. The coding and theme identification process was prepared in Atlas.ti, which is a qualitative data analysis software package that offers support for the interpretation of text (Muhr, 1991, p. 349). Categories and code names emerged from the data and quotes, keywords, or paragraphs were counted and compared followed by interpretation of the underlying context (Rosengren, 1981).

Discussion of findings Themes were aligned with research questions and categorized to identify initiatives employed to promote and execute transformation, as well as challenges experienced in the process of transformation, in a manner that addresses the research objectives. It is important to firstly discuss the interpretations of the meaning of transformation by participants before a discussion on initiatives to promote transformation and challenges experienced in transforming the mining industry. This forms part of the next discussion.

Common understanding of transformation definition It was found that the participants had varying interpretation of transformation, although common terminology regarding transformation understanding was that transformation is a cultural change but not a racial issue involving replacing white individuals by black individuals in specified positions.

Transformation was regarded by participants as a mind-set change, the act of embracing diversity, equalizing rights and creating opportunities, and doing what is right for organizations. The majority of participants acknowledged that transformation is a process that will take a long time to realize.

Research methodology

Initiatives to promote and execute transformation

A qualitative research methodology was used and data was collected through ten in-depth interviews. Participants representing mining houses listed on the Johannesburg Securities Exchange were selected by means of probability purposive sampling. This technique allowed the researchers to select and interview executives or senior managers who oversee transformation, sustainability, human resources, people management and/or employment equity for the entire company. The size of the companies ranged from small to large multinational groups.

The duration of interviews ranged from 45 minutes to one-and-a-half hours, and interviews were conducted at a place convenient for the participant. Nine out of ten interviews occurred at the offices of the participants. The overarching topic of the interview was the participant’s experience with transformation in the mining industry. This topic was supported by open-ended questions dealing with different issues related to transformation and employment equity scorecard – refer to Appendix 1 for a research questionnaire. Each participant voluntarily signed informed consent documents, and was informed about the purpose of the study and assured that their information would be treated as confidential. All interviews were recorded and were later transcribed into primary documents. Qualitative content

Key themes that emerged in this category were staff recruitment efforts, staff development, staff retention, and community development.

This study found that recruitment of HDSAs (theme 1) goes beyond traditional recruitment practice, i.e. using recruitment agencies or advertising vacant positions. As available candidates often lack specific skills and/or experience, employers interviewed for this study often offer graduate and management development programmes. The objectives of these programmes are to equip new entrants with skill sets necessary to assume specialist and management roles in the company. To recruit employees for more senior positions, all participants indicated that their recruitment drives involved rigorous headhunting efforts in search of HDSAs. Staff development is another priority in support of transformation. Indeed, all participants reported that they offer bursaries to eligible staff members to pursue studies at higher education institutions as well as on-the-job training and various external training opportunities. Although training opportunities are available to all staff members, it became clear that preference was given to HDSAs. Insightful quotes in support of this deduction were made by participants 2 and 9:

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there. And that school, the frames, steel frames came from a welding centre that we built in the community’ (Participant 8).

The findings above suggest that mining companies are committed to recruiting and developing HDSAs in an effort to drive transformation. Focus is also placed on staff retention and community development.

Challenges to transformation Unfortunately, transformation in this industry poses several challenges. The research results disclosed six themes that describe challenges to transformation – inability to recruit suitable candidates, competition for talent, mining not always viewed as a suitable career choice, cultural diversity, lack of government support, and the debate on nationalization of mines.

The most prevalent theme revealed in this category was concerns expressed by participants that they were not always able to recruit suitable candidates (supporting data obtained from nine out of 10 participants). Suitability was defined in terms of required skills, qualifications, experience, and adherence to employment equity. For example, participant 10 indicated that there is a skills shortage in regard to critical skills … ‘there is a shortage of talent particularly the black engineers. And I think statistically it has been proven that there isn’t too many of them in the country in terms of mining. I am talking mining in particular. There might be a lot of them in the system but they are not yet ready to take the position’.

Furthermore, participants agreed that qualifications and skills should be separated, skills being recognized as most desirable. … ‘when we say skills, I am not talking about someone who is going to come out with a master’s degree at [a] university and then we say that is skills. Skills come with experience, experience brings about insight.

Now that is what we are running short of’ (Participant 1).

‘I am an engineer, I know, it is an incredibly strenuous and onerous degree to get. It is not Mickey Mouse you have got to work hard. You have to have a little bit of ability, you have got to be able to think at a certain level, and you have got to be able to apply yourself, you have got to work hard. There is no free lunch in these types of skills, these critical skills and I am talking about engineering specifically’ (Participant 2). … ‘you still have got lots of vacancies in the engineering department which you can't fill today because of lack of adequate skills ’(Participant 3).

These findings suggest that competence is a combination of qualifications, skills, and experience. To meet these requirements is not easy and various perspectives were offered on whether the industry needs to lower the standards for recruitment. Participant 7 admitted that: … ‘competence is colour blind in this industry because remember that if you are at the top and you have a thousand people underground, three kilometre down there, working at the rock face, it’s got nothing to do with colour. When you are the accountable guy here, you have to pull them out.’.

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‘The government expects us to pay, spend 2% on payroll on training … we are making sure that we are focusing most of our training on core and critical skills, but we are looking specifically at the HDSA’s that they get it.’

‘We try to train people to go up, getting more core skills. So if we have got a local guy living near a mine and he comes from a local community, we teach him how to dig in the ground, we teach him how to work underground, we teach him how to do drilling, blasting, we teach him how to drive trucks … we teach him more and more’.

Fast-tracking of women seems to be a key priority in the mining industry (this notion was supported by six participants). This finding suggests that development of women and gender diversity are taken seriously. Respondents did not emphasize talent management (mentioned by only one participant), change management programmes (commented on by one participant), or diversity training as being of particular importance addressing transformational challenges.

Staff retention efforts (the third theme) were described by three participants. In order to retain staff, these participants reported that scarcity allowances were awarded to black individuals with scarce skills. All ten participants, however, questioned the validity of such allowances to retain staff.

Only two participants reported other staff retention initiatives such as home ownership schemes for mineworkers. These participants specified that their companies offered housing benefits and facilitated home ownership for miners in the form of a bond at a discounted rate.

Initiatives to promote and execute transformation, however, extended beyond staff recruitment and development and included community development projects. All participants indicated that bursaries, learnership opportunities, and training are offered to non-employees. Several examples were furthermore cited to illustrate community development programmes, such as: … ‘on the skills level, for the school leavers we have bridging schools. So we are paying for students with no bursary obligations. We are just giving them money and saying, “you are a school leaver and you are historically disadvantaged, you come from local high school somewhere in the rural areas near a mine, we give you money to do bridging for a year and then we re-evaluate how well you have progressed.” So we spend a lot on education’ (Participant 2). …‘we are providing extra lessons for Maths and Science grade 10 to 12’ (Participant 5).

Efforts protracted beyond training and all participants confirmed their commitment to community development.

Participant 7 described this commitment as: … ‘in this year alone, for example, we trained over 5000 community members on different skills programmes. Now I am not aware of any other industry that trains people outside the company other than its own. And that is typically what mining does’.

Two participants also emphasized the importance of building strong relationships with other stakeholders such as municipalities in order to implement their projects.

‘We talked to [the] municipality, they formalized it, they put in water, electricity, we have built a school

Transformation in the South African mining industry In addition to the recruitment challenges reviewed, participants indicated that recruitment of females and people living with disabilities was particularly challenging.

‘The challenge remaining on the senior management and also the demographics you will find that [for] African females in particular there is still a challenge there. But you can also realise that when we are talking about senior management we are talking about someone who has an experience of ten years or more to fifteen years’ (Participant 9).

‘We don’t have in my view a robust system or a process, whatever you want to call it, that focuses on people with disabilities and ensures that we can as a country produce people with good skills. Not just skills in terms of being a typist or as a person with a disability but professional skills amongst the people who have got disabilities’ (Participant 3).

Another recruitment challenge was sourcing suitable candidates from local communities. Although participants indicated that they try to recruit candidates from local communities, six participants reported that local candidates in mining communities are not willing to do the work that the foreigners or ‘men from the Eastern Cape or Lesotho’ are willing to do. … ‘we target people from the local communities …

What you find is they are not willing to do the same kind of work like people in the Eastern Cape and other places are willing to do’ (Participant 6).

‘They say “no it is hard thank you I am not going to I can't do it” you will find that those people would want office jobs and not go and become a rock drill operator’ (Participant 7).

Participant 6 described attitudes that local communities have towards migrant workers. … ‘that creates a lot of tension within the communities so all the people who come from the Eastern Cape or come from Mozambique or come from Lesotho, find themselves isolated because they are not accepted into the communities...In fact they are seen as foreigners in those very same communities’.

Once suitable candidates are appointed, the next challenge is to retain them. In this regard, all participants in the study confirmed that the major challenge facing mining companies is ‘a war for talent’ (theme 2) due to the inability to attract and retain talented individuals. Three participants mentioned that those who are mentored, coached, and trained are often headhunted and recruited by rivals. … ‘probably every year we get top guys, we mentor them, we coach them, we give them positions, they work, after six, 18 months they get poached by somebody else’ (Participant 1) These participants furthermore reported that they headhunt HDSAs and often pay a premium for recruiting them. One participant stated that HDSA ‘attraction bonuses’ are awarded for accepting a job with a company. However, the challenge remains that HDSAs still leave the company as soon as they find another occupation with greater financial rewards.

Further complicating recruitment and retaining staff was that positions in mining are not always a suitable career choice (theme 3). Demanding and sometimes dangerous

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working conditions are considered to be an obstacle to achieving transformation objectives in a spirit that promotes health, safety, and employee wellbeing. … ‘an issue around health and safety is a major challenge for the mines. Because sometimes the health and safety is not within the control… you get things like fall of ground… you can bar all the places and all that and then all of a sudden a rock falls out there’ (Participant 1). … ‘equipment that we are using is not [ergonomically] designed for women. We are still using lots of conventional mining. I mean that drill weighs about 24 kilos and if you are to pick it up as a woman I don’t know if I can’ (Participant 3).

Complementing the previous theme was an entrenched prevalence of stereotyping in the mining industry. … ‘it’s a very male dominated environment and I think there is still a lot of stereotype about what woman can do and what woman can’t do. So I think there is a lot of a stereotype around gender’ (Participant 6). … ‘they are entrenched here, they have been here 30, 40 years and they will tell you now you come in your 20s and you want to come with a new way of doing things.

Classic diversity challenges’ (Participant 3).

It was further reported by participants that gender and age stereotyping are linked to cultural diversity challenges (theme 4) faced in the mining environment. For example participant 6 explained that … ‘you need to understand that in our culture, the African culture, a woman is more subservient to a man.

So suddenly you are you are trying to promote woman into a supervisory position. And these men don’t understand. They say, “what the hell now a woman must come and tell me what to do?” ’.

‘I think mentoring is a big problem. I think black people get thrown into positions and then there are no safety nets. And unfortunately, when that happens and that person fails and then the argument is black people can’t do it. It’s not that. They need to be given the same support that their white counterparts were given in order to succeed’.

Participant 4, however, stressed that cultural insight is required to mentor staff: … ‘one of the barriers is the fact that you don’t have enough mentors who are like the people that you are trying to empower [HDSA] ... it is sometimes very difficult to empower people or advise people when they don’t really understand where you are coming from or when you don’t understand where they are coming from, when you can't relate... it is a human relations issue, because you must be able to relate before you mentor’.

Although participants agreed that transformation has been contextualized as the inclusion of HDSAs where they are underrepresented, it could sometimes lead to the exclusion of and discrimination against the white racial group. Such exclusion often leads to racial tension. Transformation also creates uncertainty for staff members outside the targeted race groups. Such uncertainty often results in anxiety and fear about future career prospects. Participant 2 described concerns about this challenge as follows: ‘I think more of our lower skilled white people are The Journal of The Southern African Institute of Mining and Metallurgy

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mental policies in different, in other departments of the country’ (Participant 7).

‘Where we say they are not talking to each other they talk to different [definitions], they track different things.

(Participant 8).

Participants also questioned targets and argued that meeting targets is complicated by employee behaviour and culture. One such example is improving living conditions.

From the responses received (60% of participants), it was established that the 2014 targets of improvement of housing and living conditions would not be achieved. Reasons offered for non-compliance included miners erecting shacks although housing allowances have been granted … ‘some of them get R1 800.00, then they erect these backyard [buildings], and some are renting proper rooms in the backyards of people in the nearby villages. But most of these people, 90% of them are saying, “I am here for work, I would rather build the house at home because that is where I am going to retire” (Participant 5).

But temporary residences have serious consequences ‘… but the problem is that you still have people who don’t necessarily want to live in that single room accommodation … so we may have given them that R2000 or so that is supposed to be a living out allowance, but it not being used for living out accommodation. It’s being used to supplement a secondary family that they may have in the North West while they have another family in the Eastern Cape’ (Participant 6).

Participants also questioned government’s commitment towards creating an environment supportive of transformation. … ‘it is okay for the government to say to industry, “these are the targets.” But then the government has a commensurate responsibility to make sure that they deliver the education to our young people, then we can take school leavers and educate them and take them further. We are very happy to do that. But if we have a paucity of skills and abilities in school leavers coming out, it is actually almost abdicating that responsibility to industry. And I know that the government will confess privately that they don’t have the capacity to deliver on anything that they need … so they admitted that they don’t have all the capacity to do it ‘(Participant 2).

Claims of limited service delivery by government were also major concerns: ‘One of the issues was how do we capacitate local municipalities? What role do we as industry want to play in that space? We don’t want to take over municipalities but how do we capacitate them? We’ve got a company to run but how do we capacitate them so that they can deliver because if they can deliver they take the pressure off us, you understand? For as long as they don’t deliver we will always be at the mercy of communities and we accept that we have a responsibility towards [these] communities’ (Participant 6).

‘I don’t think they have their ducks in the row … to me that’s the biggest barrier as well. Because if you say to me please implement this but you still don’t have your house in order [and] you don’t even guide me on how to do it you must forget’ (Participant 10).

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extremely afraid, they are … and that brings out a resistance … people are changing, people not wanting to take another job or scared of resigning or scared of moving on, because … they think they are not going to get opportunities’.

The outcome is often that employees who do not benefit from black economic empowerment believe that they don’t have alternative employment options, and as a result staff turnover is slow. In some instances, productivity is also affected as ‘White people are feeling so disenfranchised now.

They are feeling so reversed apartheid … they feel like second class citizens at the moment. [They] feel they have got something, but at any moment [they] can lose it and it is actually very depressing for white people’ (Participant 2).

This finding is in line with studies of Selby and Sutherland (2006) who argue that such feelings result in a breakdown of the psychological contract with existing white employees and increased loss of memory due to a lack of commitment. Racial tension also emanates when blacks feel that they are being undermined in the workplace and being unfairly criticized for the quality of their work.

Outside the control of industry participants, another challenge highlighted was the lack of support provided by government (theme 5). All 10 participants, for example, criticized the education system for not preparing learners for mining-related occupations or the standard of education provided, resulting in many learners not meeting the admission requirements for mining-related qualifications offered at a tertiary education level. All participants furthermore reported that schools in the rural areas did not have sufficient facilities and resources, such as access to a library, laboratories, and equipment for experiments. Seven participants reported that mining companies have tried to accommodate HDSA school leavers but found them illprepared to meet employers’ expectations. It was also reported that when HDSA school leavers were offered the opportunity to further their education, they often performed poorly. … ‘a lot of my students I had given bursaries to, they just couldn’t get past second year. They did first year, failed first year, did second year, failed second year.’

(Participant 2).

The current social grant system was also blamed for hindering the development of a work ethic among youth. … ‘because you will get a free house by hook or crook, you will make [the] means, … and then you will get pregnant you go to [get] social grant … That is what the youth is doing now. Even if you interview them, “I need three kids so that at least I can get this much” and then they budget already. Then when you get older you will go for the social the old age social grant. So there is always something that will be handed out. So we have created a culture of dependency’ (Participant 5).

Most participants (90%) furthermore expressed their discontent with the disjoint between different definitions in relevant Acts, frameworks, and scorecards. Accordingly, participants agreed that they were uncertain about which policies and targets apply. … ‘some policies are not aligned with other develop-

Transformation in the South African mining industry experiencing challenges in accurately monitoring mining companies’ compliance with the legislation.

‘I know that they had a challenge; they have got a challenge of monitoring companies in terms of how far they are going’ (Participant 3).

A lack of collaboration is furthermore attributed to the lack of trust and limited dialogue between the mining sector and government.

‘For as long as we treat each other with suspicion … because that unfortunately the fact, there is a lot of suspicion around the mining industry and unfortunately what happened at Marikana doesn’t paint us at a good light at all.’ (Participant 6).

The final challenge identified facing the mining industry is the industry’s uncertainty and anxiety emanating from calls for the nationalization of mines. Participants expressed different opinions about this call. Three participants opposed nationalization and claimed that insufficient information regarding nationalization exists and alleged that a political agenda was behind the call. There were also concerns about the state’s ability to run mines and questions about whether research was being done to determine the viability of nationalization if it was to become policy.

Findings This article describes specific initiatives undertaken by mining companies to transform. Most notable initiatives included staff recruitment efforts to appoint HDSAs, staff development initiatives, as well as community development.

Findings furthermore contextualized the challenges experienced by the industry participants in their quest to transform. Despite Minister Shabangu’s accusations that industry participants are not taking responsibility for the implementation of the transformation agenda, government needs to recognize that they too have a role to play and need to appreciate and assist in the current challenges experienced by the industry.

The present results are consistent with previous studies (Delloitte & Touche, 2013; DMR, 2010; Esterhuyse, 2003; Landehlani, 2008, 2013; Rungan et al. 2005; Sapa, 2010) and highlight the industry’s inability to recruit suitable HDSA candidates. Specifically, the results correlate with findings of the Landehlani Mining survey in 2008, which revealed that mining is experiencing a shortage of skills in engineering fields, technical, and artisanal skills. It was also found that ‘suitability’ is measured with regard to qualifications, skills, and experience. The combination of these elements creates competence, which is considered to be non-negotiable in this industry. This finding thus challenges Minister Shabangu’s claims that skills are available based on the number of graduates produced annually and the high unemployment rate among the educated youth. Mining houses are supporting the development of staff by means of development programmes, bursaries, and training. However, more effort is required to improve the education system so as to alleviate the skills shortages. The quality of education in South Africa, specifically in rural schools, has been identified as a barrier to transformation. The lack of skills comes from matriculants who are ill-equipped and unqualified to pursue mining qualifications at higher education institutions. Government cannot just regulate and monitor transformation progress,

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but should also be an effective agent of change by ensuring that schools are equipped with facilities and resources to function optimally, especially in key subjects such as mathematics and science. Formal collaboration agreements between local governmental bodies and mining companies could result in community development programmes and improved service delivery.

Internally, mining companies also need to review their human resources practices in support of transformation.

Special attention is required to promote representation of women and create a culture of inclusivity in the working environment. Various studies, such as Daily et al. (1999) and Erhardt et al. (2003), recognized that representation of women at senior to board level needs to be fast-tracked. Yet, more than a decade later this remains a problem in the mining industry without a justifiable reason. Focused recruitment strategies and development programmes are required to address this problem. It was furthermore found that initiatives undertaken to create transformation mostly focused on recruitment, development, and retention of HDSAs. An area that seems to be neglected is the creation of a culture of inclusivity that focuses on the entire workforce.

Inclusivity is required to create a supportive and caring working environment that embraces diversity and change.

Human resource initiatives are thus required to empower employees to embrace change and diversity. Change and diversity programmes could, for example, lower existing tensions among different racial and ethnic groups. It is furthermore suggested that mentorship programmes are introduced for HDSAs that are structured, inclusive, and unbiased.

Results of the present research furthermore uncovered challenges that are often beyond the control of mining companies. Supporting findings from the DMR (2009), it was found that tension exists between local and migrant labour.

In terms of recruitment, it was found that applicants from local communities were often not willing to assume labourintensive or lower skills jobs in mining. Appointment of migrant workers also created various social problems that could increase tension in the workplace. The current research furthermore confirmed the findings of Limpitlaw et al.

(2005), that employment in the mining industry is not always considered a suitable career choice due to adverse working conditions. Participants emphasized that health and safety conditions are at times challenging, female employees are often subjected to gender stereotyping, and that the machinery and equipment used underground are still conventional and not designed for women.

Initiatives undertaken by mining houses are furthermore compromised by employees’ own choices and behaviour. For example, when mining companies build houses for mineworkers near the mining operations to conform to government requirements, mineworkers often elect not to make use of such schemes and rather use the allowance to obtain their own accommodation. In many instances, alternative accommodation arrangements do not comply with government specifications and mining companies are blamed for poor living conditions. It was also found that some mineworkers have second families and are at times subjected to garnishee orders or experience poor living conditions because they are unable to support two families. Finally, The Journal of The Southern African Institute of Mining and Metallurgy

Transformation in the South African mining industry industry participants are subjected to a broad legal framework with disparate definitions and targets. This framework, as well as uncertainties created by calls for nationalization, often leads to a lack of trust between key stakeholders.

Conclusions and recommendations This research recognizes the transformation initiatives undertaken by mining companies, and challenges the assumption that mining companies are reluctant to transform. Although the industry has its laggards, the results from this study suggest that mining companies have boughtin and committed to transformation. Nonetheless, the complexity associated with transforming the mining industry requires effective collaboration between government and industry. Open dialogue and trust are key requirements for addressing current challenges and creating solutions to benefit not only individual stakeholders but the country as a whole. It is also imperative address competency development beyond schooling for HDSAs to improve their knowledge, skills, and experience, which are vital for the mining industry. However a balance on demographic transformation should be created and racial tensions could be avoided. This will ensure compliance with human rights as prescribed in the South African Constitution on equal opportunities for all.

The study recommends that all stakeholders should agree on the definition of transformation in the South African mining industry. There is also a strong need for a common transformation implementation policy that clearly outlines the strategy, objectives, and targets for transformation.

7.

6.1 What are the major challenges facing your company in terms of EE scorecard implementation, submitting EE plans and reports?

6.2 What are the barriers to transformation in your company?

6.3 What initiatives are currently undertaken to address these challenges?

6.4 What is the way forward?

• If no, answer the following questions: 6.5 What are the successes recorded in your company?

6.6 Highlight initiatives undertaken?

6.7 What were the challenges faced?

6.8 What is the way forward?

Comment on the nationalisation of mines debate Describe your future plans for your company in addressing EE scorecard targets?

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2.

What does your company understand about the Mining Charter and the Mining Scorecard?

3.

Who are the beneficiaries of the Mining scorecard?

4.

What does your company understand by the term Historically disadvantaged persons (HDSA’s) 5.

Do you have an employment equity plan for this year or reports for the past 5 years?

6.

The mining industry is currently faced by charges of the lack of transformation, is this a case with your company?

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What is your position in this company?

2.

How long have you been with this company?

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What is your functional and educational background?

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http://dx.doi.org/10.17159/2411-9717/2015/v115n8a3

Reconciliation along the mining value chain

by A.S. Macfarlane*

This leads to the notion of having a system in place for accounting for the metal. The University of Cape Town defines metal accounting as ‘… the estimation of (saleable) metal in a mine and subsequent process streams over a defined time period.

Comparisons of estimates, from different sources over a specific time period, are called reconciliation.’

Traditionally, metal accounting has consisted of geological reconciliations of mineral resource to mineral reserve conversion, mine planning reconciliation related to the reconciliation of long-term plans to short-term plans, grade control reconciliation of head grades to sampled or block grades, and mine survey reconciliation of tonnage discrepancies and mine call factors. In addition, metallurgical reconciliation has tended to be based on metal balancing, followed by some commercial metal accounting that links dispatched product to sales quality, quantities, and revenues. Miners disparagingly refer to this plant process as ‘millmatics’.

These reconciliations are all done for the particular departmental purpose at hand, and have been done to varying levels of quality and effectiveness, but seldom integrated into an end-to-end process.

Recognizing this, AMIRA introduced the concept of an end-to-end metal accounting system through a series of global workshops in 2001, which culminated in the formation of the Metal Accounting Project P754 in 2003. As a result of global research work, the Code for Metal Accounting was published in its current version in 2007 (Morrison, Gaylard, 2008).

Synopsis Metal accounting and reconciliation is an increasingly important governance issue in all mining operations, in that it is required, from a risk management perspective, that the company is in control of its product throughout the whole mining value chain.

Reconciliation is a grossly misunderstood term. It means different things to different people, and therefore one of the purposes of this paper is to ensure a holistic and integrated understanding of ‘reconciliation’.

Previously, where reconciliation was done, it was often for internal control purposes and loosely applied, sometimes with a low degree of confidence and understanding of the underlying parameters and their natural variability.

In most mining operations, issues such as dilution, stope performance, and recovery are critical to profitability and long-term value, and thus understanding and control of these key value drivers is essential, not only from the governance perspective, but also from the perspective of maximizing shareholder returns.

In order to implement a system for metal accounting and reconciliation, it is important that the cause and effect of these value drivers is understood, and that a systematic control system be established.

While a number of off-the-shelf solutions exist for this work, it is the contention of this paper that a full understanding of the metal flow, its variability, its underlying loss potential, and its control points is necessary before a systematic approach to reconciliation can be undertaken.

The paper also advocates that this approach should ensure that the reconciliation system clearly addresses the reconciliation needs, within a consistent framework. Such a framework has been developed by AMIRA, in terms of a metal accounting code. However, up until now, this code has been aimed at plant processes, whereas this paper shows how the concept can be extended to cover the full reconciliation requirements for a base metal mine.

Keywords metal accounting, reconciliation, AMIRA, grade control.

Introduction

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* Tomahee Consulting. © The Southern African Institute of Mining and Metallurgy, 2013. ISSN 2225-6253. This paper was first presented at the, Optimisation of mine value chain from resource to market Conference 2013, 7–9 May 2013, CSIR Convention Centre, Pretoria.

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Reconciliation is defined as ‘the process of finding a way to make two different ideas, facts, etc., exist or be true at the same time’. In terms of a mining process, this refers to the comparison of measures and estimates along the value chain, and at different points in time, in order to track and optimize metal recovery.

Reconciliation along the mining value chain During the process of development of the Code, the University of the Witwatersrand was involved in an attempt to incorporate the geological and mining components into the study, in order to produce an end-to-end accounting system.

Unfortunately, at the time, limited support for this was forthcoming from mining companies, with the result that the research and the Code itself have tended to concentrate on the mineral processing part of the value chain only. Hence, the traditional stumbling block, the reconciliation between the plant metal balancing system and the mine accounting system at the point of delivery from the mine to the plant, remains in place.

This paper attempts to show that the AMIRA Code principles and process can indeed be extended to the full value chain as an end-to-end reconciliation system.

Reconciliation as a governance issue Corporate governance is no longer an issue associated solely with boardroom responsibility and fiduciary duty. Instead, it requires that corporate level governance and risk management is cascaded into the tactical and operational levels of the company, at the value chain and professional support levels respectively. This is supported and endorsed by the following: ➤ Sarbannes-Oxley requirements for risk management and reporting ➤ King III requirements for integrated reporting, which require integrated risk reporting ➤ Impending development of the ISO 31000 standard on risk management ➤ Increasing shareholder demands for control and optimization of product and earnings ➤ Moves towards more disclosure in terms of reporting of mineral resources and mineral reserves, and their valuation ➤ Reconciliation and metal accounting as essential components of effective mineral asset management (management of the principal asset of the company) ➤ Requirements for optimization of products where logistical or competitive advantage constraints exist.

The reason for metal accounting along the mining value chain A diagrammatic representation of the value chain for a typical base metal operation producing copper and cobalt is shown in Figure 1 (Macfarlane, 2011).

The reasons for adopting an end–to-end approach to metal accounting along the value chain can be summarized as follows: ➤ To ensure that the value chain provides the basis for metal accounting ➤ To enable reconciliation of product mass between successive and relevant points along the value chain ➤ To ‘protect’ grade and product from loss along the value chain ➤ To ensure the mineral resource inventory is effectively converted into saleable product ➤ To identify acceptable tolerances and ranges of variability along the value chain ➤ To ensure that practical targets for ‘modifying factors’ such as dilution and recovery are identified through measurement and analysis ➤ To ensure that these factors are monitored and controlled within acceptable ranges ➤ To ensure that competence and control are exhibited on these critical variables ➤ To ensure reliable reporting of these variables in public reports ➤ To include metal accounting and reconciliation as essential components of the MRM and enterprise risk management system ➤ To ensure integration of reconciliation and metal accounting along the value chain.

Within the mining environment, specific and unique reasons exist that provide further compelling reasons for integrated reconciliation systems to be developed and implemented. These include the following. ➤ Base metal deposits vary internally in terms of their characteristics, often resulting in mineralogical changes

Figure 1—Base metal mining value chain

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➤

➤

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➤

➤ ➤ ➤ ➤ ➤

as mining progresses through the orebody, from oxides through to sulphides, and through other mineralogical geozones Downstream processes are often seen as the valuecreating parts of the chain. If this is the case, then at a minimum the plant needs a consistent feed in terms of volume, hardness, quality, size, and grade in order to perform to design. Alternatively, it should be recognized that value can be added upstream, where the total chain can be optimized, through a complete accounting and optimization process The array of different mining methods results in different challenges in terms of grade control and tonnage reconciliation. In most massive mining operations, grade control programmes must be specified in terms of protocols that define sampling points and methods, appropriate for the method, from the creation of the grade/block model to the final product dispatch. Grade variability in such operations is often high, and this variability must be understood from various phases of delineation drilling. Sampling methods and locations must then be determined, whether this be on the basis of sampling theory and semivariogram analysis to define sampling methods and densities, or practical considerations of the practicality of drawpoint sampling, and sample turnaround time. In all these instances, reconciliation of grade and tonnage to the block model and sampling data is critical for operational control Dilution control is particularly critical, as shown in Figure 2, which indicates the impact of dilution on the net present value (NPV) of a sublevel open stoping gold operation. In such operations, although planned dilution (through design) may be 10%, dilution in individual stopes of up to 70% is possible and not too uncommon. Reconciliation is essential in order to identify and control such dilution impacts Variance in grade and tonnage from the grade/block model or the ongoing face sampling results must be investigated in order to assess whether some geological or mineralogical change has occurred, or some change in draw control or mining sequence has resulted in a negative impact. This can be done only on the basis of sufficient information being available through a reconciliation system In platinum mining, strict control of product and mining mix is essential to improve PGE recoveries In gold mining, issues resulting in a poor mine call factor can be identified only through effective reconciliation systems In iron and manganese, optimization of dispatched grade and size can be managed only if a complete mine–to-market reconciliation system is in place In diamond mining, diamond recovery can be improved through reconciliation and optimization systems that control geometallurgical and liberation effectiveness In coal mining, control of value along the value chain is essential through effective control of calorific value, water and ash content, and discard percentage.

iation system, a systematic approach is necessary. This is because of the multivariate nature of the problem, the complexities of the system, and the multidisciplinary approach necessary for full and integrated reconciliation.

The AMIRA framework provides such a system, and can be adapted reasonably easily as the basis for the system.

The main topic areas and sequence of such an exercise would be as follows, relying on the AMIRA Code principles.

Define the purpose and objectives of the reconciliation exercise This requires the determination of the strategic, governance, and operational purpose, and the scope of the requisite programme. In other words, what is the reason for implementing the system, and what are the reporting requirements of the system? This is a strategic planning exercise that can be conducted using a multidisciplinary strategic workshopping approach to define the reconciliation charter.

Within this phase, it will be necessary to define the reconciliation boundaries that are appropriate for the operation at hand. Such reconciliation boundaries may be: ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤

Block model to sampling reconciliation Sampling to broken reconciliation Block model to broken grade reconciliation Plan to actual broken grade and tonnage reconciliation Plan to actual mining volume broken Ore broken tons and grade reconciled to ore tons and grade hoisted or trammed Hoisted ore tons and grade reconciled to plant feed head grade and tonnage Stockpiling tonnage movement reconciliation Plant balance reconciliation Plant to product reconciliation Dispatch to customer reconciliation (includes all logistical inventories, stockpiles etc.)

Model to mine reconciliations Mine to plant reconciliations Plant to product reconciliations Resource to reserve reconciliation Period-on-period reconciliations.

Once these are defined, a picture such as the one in Figure 3 can be created, that indicates possible reconciliation arcs along the value chain.

Adopting a systematic approach to reconciliation

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In order to develop and implement a full end-to-end reconcil-

Reconciliation along the mining value chain

Figure 3—Defining reconciliation arcs along the value chain

Broadly, these reconciliations are subdivided into reconciliations that are required for management control, which are from the model to the plant delivery, and reconciliations that are required for commercial reconciliation between the plant dispatch and the customer. These are required in terms of commercial agreements and contracts, and in case of dispute and litigation.

Define all ore and product flows, now and in the future This requires that a full and detailed process map from end to end be developed, that is from the geological model to the final product. Usually, these exist in some form of diagrammatic flow sheet, with detailed flow sheets in the mineral processing circuits. However, the mining flow sheet is usually loosely defined, and does not take account of illicit stockpiling possibilities, ore loss routes, and lock-up tonnages. Additionally, all opportunities for cross-tramming, cross-hoisting, and surface ore movements must be taken into account.

Furthermore, projects that are under development often show ore flows that are in transition, being temporary arrangements to be modified or changed later. Thus, ‘as-is’, as well as future, ore flow maps need to be developed.

Once a process map that shows all flows for each phase of the project has been developed, the next part of the process is to define on this map all current and future measuring points, and all transport arcs.

Conduct a full risk assessment A risk assessment should be conducted on the current system and the process flow map, highlighting all the following risks.

Once the mapping exercise has been completed, this can be done on the basis of either a fault tree analysis, or a causeand- effect analysis. In the case of the fault tree analysis, the ore flow needs to be tested in terms of defining the risks of

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poor or inaccurate measurement at measuring points, or the risk of ore or metal loss along transport routes.

Thus, in each of these areas, the risks must be identified and quantified in terms of the following: ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤ ➤

Weighing, sampling, and analytical risks Weighing equipment, and methodology risks Possibility of incorrect recording of data Possibility/risks of inaccurate sampling Risks on sample security Risks on estimates of stockpiles Risk of arbitrary methods used to assess/calculate metal recoveries Risks of inclusion of illegitimate/biased data Risks of incorrect calculations Risks of real loss of product during transport Risks as a result of incompetency of operators.

Risks should be assessed on the basis that ore and waste flows through the value chain, along transport arcs, and between measuring points. Thus the fundamental questions to be asked of each arc is ‘what is the likelihood of loss along the arc?’ and at the measuring points, ‘what is the likelihood of poor or inaccurate measurement?’ A systematic approach, using process experts, will answer these questions, recognizing the tolerance/variability that is expected from the system.

Integrity of data Once the ore flow and its risks have been identified, it is then necessary to define all data requirements and data capture systems and technologies throughout the end-to-end process.

Data variability and tolerance levels must also be defined. In the case of sampling information and data, variability is defined from the sampling data itself. On the assumption that the sampling density and methods are appropriate for the variability of the deposit (by no means a sure assumption), The Journal of The Southern African Institute of Mining and Metallurgy

Reconciliation along the mining value chain then the samples themselves can be used to define the natural variability of predicted grade, and the definition of short-range grade models for short-term planning processes.

Analysis of data can be done only on real-time data.

Monthly averaging is of no use: data on critical variables must be based on real-time data analysis and trends and variability, to assess the natural ‘noise’ in critical variables.

Analysis of data, whether on the operation concerned or a similar operation, should allow an assessment to be made of the variability of the data. In the case of the mine call factor illustrated in Figure 4, this would allow assessment of the variability in terms of a statistical function, such as standard deviation or coefficient of variability, and the definition of 90% probability limits of the expected range of outcomes.

An example of such an analysis, based on the data in Figure 4, is illustrated in Figure 5, which shows the 90% confidence limits on the variability of the mine call factor.

The example chosen illustrates an extreme case, which is actually a diamond operation.

Assess mass measurement The next step is to define mass measurement requirements and mass balancing in terms of: ➤ Accuracy, precision, and methods ➤ Mine tonnage assessments from all sources

➤ ➤ ➤ ➤ ➤

Mass flow measurement systems and calibration Assessment of all inventories and mass in process Assessment of relative density and moisture contents Assessments of fill factors and loading Definition of mass measuring points, technologies, and accuracies ➤ Mass balancing between points and over the whole process.

This part of the process defines the measuring points and technologies that should be implemented, based on critical control points and variability. This must include stockpiling operations, whether these are planned or illicit stockpiling that may be done to protect departmental interests, such as bonus payment. All these possible sources of tonnage discrepancy should have been identified in the process map.

Often, mass measurement technologies are not perceived as being robust enough to deal with mining conditions, but this is usually a preconceived notion that is aimed at avoiding the tight control that metal accounting and reconciliation systems require. For example, underground mechanized mining operations can easily provide greater precision on mass balancing, through the use of cavity monitoring systems, bucket and bowl weight indicators, and vehicle tracking and fleet management systems, all linked through to a full tonnage control system that measures and reports in real time.

Such a system was successfully introduced at Finsch Mine, where full real-time fleet management and dispatch systems were introduced, to the extent of developing an architecture to link mine planning, draw control, dispatch, and fleet management systems into an integrated management information system, driven through a central database.

Sampling In terms of sampling, the following must be defined with respect to grade sampling:

Figure 5—Statistical analysis of mine call factor variability The Journal of The Southern African Institute of Mining and Metallurgy

➤ Appropriate sampling points, at the face and at points along the process flow (Figure 6) ➤ Sampling methods appropriate for the orebody at hand ➤ Sampling quality assurance and quality control ➤ Appropriate and reliable sampling technologies ➤ Sampling accuracy ➤ Sampling precision ➤ Sample storage, transport, supervision, and security ➤ Assay quality assurance.

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Figure 4—Mine call factor trace for a diamond mine

Reconciliation along the mining value chain Sample preparation and analysis This includes the definition of sample splitting and assaying protocols and quality assurance, through the use of validation, checks, duplicate samples, standard samples, blank samples etc.

All of these considerations should be included in a grade control policy and protocol document that provides a governance policy on grade control to be defined.

The grade control protocol should include: ➤ Data collection – Analysis methods – Modelling procedures – Planning for data collection ➤ Definition – Cut-off grade application – Level of selectivity – Link and reconciliation to planning ➤ Monitoring and control – Develop grade control plan – Staking and delineation – Monitoring of loading – Load dispatch control – Blending strategy – Stockpiling strategy – ROM sampling procedures ➤ Stockpiling allocation/definition – Monitoring/control of stockpiles – Truck/vehicle count – Measurements ➤ Reconciliations – Truck factors – Reconciliation of model to actual to stockpile.

Stocktakes and inventories It is necessary to define methods and frequencies for all stocktakes, stockpile measurements, inventory measurements, and work in progress tons and grade. This must be active, in that the regularity of measurement needs to be assessed. Month-end reconciliation is not enough: realtime measurement of all stockpile measurements and reconciliations must be undertaken. This can be achieved through weighline control.

Metal balancing Metal balancing boundaries and control points must be identified, in terms of mine outputs, representation/process and measurement bases, to include all factors, boundaries, recoveries, reconciliation requirements, and calculations and controls. This is throughout the whole length of the value chain.

Reporting Data collection systems and information reporting and dissemination in terms of management controls, KPIs, and reporting must be established, as well as variability analysis, so that proactive control at the right level can be established.

Reporting must be based on the critical success factors that drive the process, and on the trends and ranges of the measured variables. This takes account of the natural variability that exists, and allows reporting on an exception

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basis when out–of-range trends start to emerge. This, however, is still based on continuous real-time measurement.

Furthermore, reporting must be defined at the right level, where reporting information is matched to organizational responsibilities and key performance indicators.

As far as possible, measurement should focus on leading indicators that show that a trend is emerging, so that proactive intervention and correction can be effected.

Responsibilities Responsibility and accountability boundaries must be established so that performance management systems can be developed and implemented. AMIRA requires the appointment of a Competent Person for metal accounting, and this principle should be adopted whereby an overall Competent Person takes end-to-end responsibility for metal accounting and reconciliation, while individual persons in particular areas of expertise, form the Competent team.

The overall Competent Person should maintain an air of independence, and therefore it is logical for the mine surveyor to fulfill this role within the MRM organization.

Clearly, this does not mean that this person takes control of the numbers in, say, the plant. However, the person should have access to the plant, in order to collect and integrate the metallurgical numbers that have been derived by the metallurgist.

Accounting and auditing The system developed and implemented must be auditable and transparent. Thus, data integrity is essential, and audit trails must be available to indicate and track metal flow along the value chain.

External audit and validation is desirable.

Data, controls, and systems must be linked to enterprise accounting and auditing systems to ensure best practice compliance on an ongoing basis. This will then form part of the company risk management process and governance procedures.

This is important where poor recoveries, dilution, or loss are occurring, in that investigative processes to rectify these discrepancies must be based on sound data and information.

This may be done using a cause-and-effect approach as outlined in Figure 7, as applied to a dilution problem.

Once data is available, quantitative investigations can be conducted on real data, and sustainable improvement initiatives can be put in place.

Further analysis can then be conducted to assess the real and apparent causes of metal loss (Figure 8), using the kind of fault tree analysis referred to earlier in the paper.

Optimization Only once all of the above is in place, and the system is stable and under control, can true mineral resource throughput optimization opportunities be realized.

Base metal mining operations in particular usually exhibit opportunities for mine-to-mill optimization, whereby stable and consistent plant performance can be achieved. This may, for example, be in relation to ore feed size and mineralogy, resultant energy consumption in the mills, and acid consumption in the leach circuits. However, this balance can The Journal of The Southern African Institute of Mining and Metallurgy

Reconciliation along the mining value chain

Figure 7—Cause and effect analysis of dilution

1. Understand the problem • Define extent of the problem • Define oreflows • Analyse date: use statistical analysis • Conduct Fault Tree analysis workshop • Conduct mineralogical studies • Literature search

2. Investigate the problem

Real • Conduct liberation studies • Mass sampling • Underground observations • Fragmentation studies • Asset protection

Apparent • Investigate sampling size, frequency density, methods • Assaying QC • QA/QC on sampling and assaying • Check Geomodel (dilution, fault loss) • Assess variability

3. Establish sustainable control • Define measuring points • Establish measuring systems • Apply analytical software to data collection • Establish control systems and procedures

Figure 8—Real and apparent loss

Conclusions Full and complete metal accounting is essential for control and governance along the value chain. In order to establish such a system, a systematic approach that is common along the full value chain should be adopted. The AMIRA Code for Metal Accounting gives good guidance on how this should be achieved, in that it allows a process that includes the identification of practices, processes, people, and technologies that will allow such a system to be established.

Clear accountability must be established for the management of metal accounting and reconciliation, as an essential component of the mineral resource management The Journal of The Southern African Institute of Mining and Metallurgy

function on the mine.

Only once a full, integrated system is in place, that captures and analyses information on a real time basis, can investigative work and optimization studies be put in place.

Thus metal accounting and reconciliation provide an essential foundation for real value creation.

References MORRISON, R.D. and GAYLARD, P.G. 2008. Applying the AMIRA P754 code of practice for metal accounting. MetPlant 2008, Metallurgical Plant Design and Operating Strategies, Perth, WA, 18–19 August 2008. pp. 3–17.

MACFARLANE, A.S. 2011. Metal accounting along the base metal value chain.

Sixth Southern African Base Metals Conference, Phalaborwa, South Africa, 18-21 July 2011. The Southern African Institute of Mining and Metallurgy. pp. 429–442.

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be achieved only once the full, quantitative ore flow control system is in place.

The Southern African Institute of Mining and Metallurgy in collaboration with the South African National Institute of Rock Engineering are organising the

INTERNATIONAL SYMPOSIUM ON SLOPE STABILITY IN OPEN PIT MINING AND CIVIL ENGINEERING

SLOPE STABILITY 2015 12–14 OCTOBER 2015 In association with the

15–16 OCTOBER 2015 Cape Town Convention Centre, Cape Town BACKGROUND pen pit mines are being planned to increasingly great depths, often beyond the current experience and knowledge base, with the exception perhaps of some very steep and high natural rock slopes. Despite recent advances, there remains significant challenges in understanding the mechanisms of slope behaviour and failure, and methods of stability analysis for such slopes. In an attempt to bring together the state of the art capabilities in these fields, as well as new research and developments, the South African Institute of Mining and Metallurgy (SAIMM) in collaboration with the South African National Institute of Rock Engineering (SANIRE) are organising a specialist international symposium on the Stability of Rock Slopes, to be held in Cape Town in October 2015. Cape Town is considered to be a very appropriate location for the symposium. There is a wide choice of hotels, within a few minutes walk of the symposium venue. In the area there are many steep natural rock slopes, and there are significant hard rock quarries within about an hour’s drive. Technical site visits to various natural slopes and the quarries will be arranged.

O

EXHIBITION/SPONSORSHIP Sponsorship opportunities are available.

Companies wishing to sponsor or exhibit should contact the Conference Co-ordinator.

WHO SHOULD ATTEND The symposium will address important developments in design, analysis, excavation and management of rock slopes. The organising committee will aim to provide a programme containing valuable and stimulating information. This should be of interest to: > Geotechnical - rock engineers and engineering geologists > All open pit mine personnel > Mine management > Consulting engineers > Road and rail engineers > Contractors > Blasting specialists > Research and academic personnel > Hydrogeologists > Manufacturers and suppliers of slope stabilisation equipment > Slope monitoring and survey specialists > Manufacturers and suppliers of slope monitoring and survey equipment and geotechnical instrumentation > Government minerals and energy personnel > Health and safety personnel and officials.

For further information contact: Head of Conferencing, Raymond van der Berg SAIMM, P O Box 61127, Marshalltown 2107 Tel: (011) 834-1273/7 Fax: (011) 833-8156 or (011) 838-5923 E-mail: raymond@saimm.co.za Website: http://www.saimm.co.za

Keynote Sp eakers G. Beale L. Lorig D. Stead

OBJECTIVES > To document information regarding state of the art rock slope design and excavation > to gather together the latest information regarding the application of probabilistic design and risk analyses of slopes > to provide information on state of the art slope monitoring and maintenance techniques > to present valuable, illustrative and interesting case studies.

Sponsors:

http://dx.doi.org/10.17159/2411-9717/2015/v115n8a4

Effects on entrainment of serpentines by hydrophobic flocs of ultra-fine copper-nickel sulphides during flotation

by M. Tang*, S. Wen*, and X. Tong*

Slime coating is one of the most common ways for serpentines to contaminate metallic mineral concentrates during traditional flotation of coarse sulphide particles. This could pose quite a complicated and challenging problem in the case of some types of low-grade and finely disseminated Cu-Ni ores bearing high serpentine contents. This is the case for the copper and nickel sulphides from the Yunnan Mine, China. Previous batch flotation tests of this ore resulted in satisfactory recoveries of 86.92% Cu, 54.92% Ni, and 74.73% Pt+Pd, and concentrate grades of 4.02% Cu, 3.24% Ni, and 76.61 g/t Pt+Pd. However, the MgO content in these concentrates was more than 19%. In the current study, microflotation tests and settling rate tests were introduced to investigate the effects of a combination of strong collectors (a 2:1 weight ratio of butyl xanthate and butyl ammonium dithophosphate) on entrainment of serpentines in metallic mineral concentrate, as well as visual observations of the concentrates in suspension using still photography. All test results indicated the presence of serpentines entrapped in the hydrophobic flocs that resulted from these collectors, even with the use of effective gangue depressants. These strong collectors are used to flocculate the ultra-fine sulphides by forming loose and ‘fluffy’ hydrophobic flocs. However, these hydrophobic flocs may also be able to load or entrap some serpentine slimes into the concentrate, and this entrained serpentine could be harder to remove by using depressants or intensified conditioning than serpentine slime coating on the particle surfaces.

Keywords sulphide flotation, serpentine, hydrophobicity, ultra-fine, entrainment.

Introduction Serpentine minerals, which have the generalized composition (Mg,Fe)3Si2O5(OH)4, can be easy to crush and grind due to their convoluted and bent layered structures. These serpentines can hinder the enrichment of some metallic minerals and dilute their concentrates by entrainment during flotation. The resulting high pulp viscosities, slime coatings, and high content of dissolved ions adversely affect the flotation recoveries of copper and nickel sulphides, and high levels of MgO entrained in the concentrates lead to heavy penalties from smelters. Since the quality of nickel flotation concentrate depends heavily on its MgO content (generally less than 6–7% for No.1 or 2 grade nickel concentrate according to the standard requirement from National Nonferrous Industry), reducing the entrainment of serpentine during flotation is becoming more and more pressing. Previous research has focused on the collectors and depressants that The Journal of The Southern African Institute of Mining and Metallurgy

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* Department of Mineral Processing, Kunming University of Science and Technology, China © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Dec. 2013 and revised paper received Nov. 2014.

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Synopsis

are commonly used in copper-nickel flotation, as well as their interaction mechanisms. The traditional collectors, like xanthate (Senior et al., 2005; Malysiak et al., 2002; Bozkurt et al., 1998), dithiophosphate (Wiese et al., 2006), or mixed/combined collectors (Glembotaskii, 1958; Lotter and Bradshaw, 2010; Wiese et al., 2008) are used to enhance the hydrophobicity of copper or nickel sulphides during flotation. Lu et al. (2011) indicated that potassium amyl xanthate is adsorbed on a mixture of pyrite and serpentine, but that adsorption is minimal on serpentine alone in alkaline pulp. Feng et al. (2012) found that the longer the hydrocarbon chain of xanthate, the lower the Ni grade and recovery, indicating that serpentine entrainment is influenced significantly by the type of collector. These collectors might not affect the natural floatability of serpentine directly, as indicated by Edwards et al. (1980). However, they may possibly influence the entrapment of serpentine indirectly by forming hydrophobic flocs of ultra-fine metal values. Also, slime coating is a common way for serpentines to be incorporated into metallic mineral concentrates. However, according to some researchers (Wellhame et al., 1992; Bulatovic, 1999; Chen et al., 1999), slime coatings can be reduced or removed from nickel mineral surfaces through strong and lengthy agitation. A possible additional mechanism for serpentine entrainment during the flotation of this type of ultra-fine ore, other than water recovery or slime coating, is indicated by some unsatisfactory results, coupled with high content of MgO in concentrates, even when using efficient depressants like carboxymethyl cellulose (CMC) and guar gum combined with high-intensity agitation (Wellhame et al., 1992; Senior et al., 1994; Senior and Thomas, 2005). The flocculation behaviour of ultra-fine

Effects on entrainment of serpentines by hydrophobic flocs of ultra-fine copper-nickel sulphides during flotation metallic values during flotation may have an important influence on the entrainment of serpentine.

Although the mechanism of serpentine entrainment into sulphide concentrates during coarse particle flotation by slime coating or water recovery is well established, the possible effects of hydrophobic flocs of fine sulphide particles resulting from strong collectors is not sufficiently understood.

In this study, batch flotation tests, microflotation tests, and settling rate measurements were conducted with the object of investigating the indirect effects of strong combined collectors on the entrainment of serpentines during flotation of finegrained copper-nickel ore containing Pt and Pd from the Yunnan Mine, China, with particular emphasis on the role of the hydrophobic flocs of ultra-fine metallic sulphides resulted from the use of these collectors.

from Zhuzhou Mining Equipment, China) for 10 minutes to a size of 80% -74 µm. The flotation tests were conducted in a mechanical laboratory machine using a 1.5 L cell (XF-D) for the roughing and scavenging stages and a 0.5 L cell for cleaning. A mixture of BX and BA in a 1:2 ratio was used as the main collector. Sodium silicate was employed as dispersant of gangue minerals, and pine oil as frother. All the tests were conducted at natural pH (6.7). The rougher tails were re-ground, and cleaning and scavenging were conducted in the 0.5 L and 1.5 L cells for 5 and 8 minutes respectively.

Suspensions of the final concentrates from the above tests were photographed at high resolution using an Olympus camera. The concentrate, as well as mixtures of middlings and tailing from the bench flotation tests, was filtered, dried, weighed, and analysed for Ni, Cu, and PGMs.

Microflotation tests and visual observation

Experimental Materials The finely disseminated ore from Yunnan Mine, in China, assays about 0.15% Cu and 0.21%Ni, with minor amounts of Pt and Pd. The main sulphide minerals are chalcopyrite, penlandite, pyrrhotite, and violarite. The gangue minerals are dominated by serpentines, which account for 75% of the MgO. The platinum group metals (PGMs) in Yunnan ore are strongly associated with the copper-nickel sulphides, especially nickel sulphides.

Samples of pure pentlandite and serpentine (>99%, -20 µm) were purchased from Jinchuang Mine, China. The flotation reagents included butyl xanthate (BX), butyl ammonium dithophosphate (BA), sodium silicate, (carboxymethyl cellulose, CMC), and pine oil, which were purchased from Kunming Metallurgical Research Institute, Yunnan province, China, and KN and OC (our patented reagents), which are traditional reagents that have been modified or combined with other reagents. OC acts as an inorganic activator that can effectively activate oxidized pentlandite; its major component is copper sulphate, accompanied by other salts. KN is modified guar gum, which acts as a depressant in combination with CMC. The open-circuit flow sheet and experimental conditions for the batch flotation tests and microflotation tests are shown in Table I and Figure 1.

Methods Batch flotation tests To prepare the sample for the batch flotation tests, 500 g of the sample was ground in a ball mill (XMQ-67Φ240×90 mm,

A Hallimond tube was used for microflotation of pure pentlandite and serpentine. Measured weights of the pure minerals were washed with dilute acid, dried in a nitrogen atmosphere, then conditioned by the combined collectors (BX plus BA in a 1:2 ratio) at a chosen dosage, and floated at a suitable bubble size for 1 minute. The concentrate was collected, dried, and weighed, and suspensions photographed as for the batch flotation products.

Settling rate measurements To prepare the samples for settling rate measurement, a head feed was ground to a size range of 100% passing 40 µm,

Table I

Conditions for the bench-scale flotation and microflotation tests Experimental conditions Constants Cell type Cell volume Solid feed mass Impeller speed Air flow-rate Ore Feed size Collectors (type and dosage) Concentrates collected Depressant types

XF-D 1.5 L, 0.5 L 0.5 kg 800 rpm 4 L/min Yunnan mine 90 % passing 37μm BX (100 g/ton), DA (50 g/t) 8 min CMC + KN

Figure 1 – Schematic flow sheet of the batch flotation tests

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Effects on entrainment of serpentines by hydrophobic flocs of ultra-fine copper-nickel sulphides during flotation then dried and sampled. A concentrate sample from batch flotation tests was treated with Na2S as de-collector in boiling water for 1 hour, then filtered, dried, and sampled. A 150 g/t addition of BX plus BA in a weight ratio of 2:1 was made to both samples, which were then stirred for 3 minutes in the beakers before settling rate measurement. The settling rate tests were performed using a 250 mL graduated glass cylinder (280 mm high and 40 mm in diameter).

Results and discussion Effects of combined collectors on entrainment of serpentines in batch flotation tests The effects of dosage of the combined collectors (BX plus BA at 1:2 weight ratio) on nickel recovery and grade in concentrate from batch flotation tests are shown in Figure 2.

A significant increase in Ni recovery to the concentrate (10% increase) was obtained by increasing the combined collector dosage from 90 g/t to 150 g/t, but at the cost of a sharp drop in the concentrate grade, from 0.92% Ni to 0.70% Ni. According to previous research (Bulatovic, 2003), strong collectors like xanthate can efficiently and selectively recover fine metallic sulphides, but there may be little effect on serpentines even at increased dosages. The much more hydrophobic flocs of the ultra-fine metal minerals resulting from higher dosages of these combined collectors may lead to a significant improvement in the recovery of Ni to concentrate, but their ‘fluffy’ structures may cause more serpentine slimes to be trapped in the concentrate, even when using the efficient depressants (CMC plus KN). It can also be seen from Figure 2 that both the recovery and grade of Ni in the concentrates deteriorated at combined collector additions of more than 150 g/t, indicating that increased collector dosages do not form more hydrophobic flocs, due to the limited number of the free-liberated metal minerals.

large and fluffy hydrophobic flocs, thus substantially increasing their settling rates. However, much less serpentine was observed in supernatants from samples digested with BX plus BA than those without any collectors after the first 8 minutes. This indicates that using these collectors may promote the entrainment of serpentine due to the formation of many fluffy hydrophobic flocs of metal minerals as ‘containers’, which is in accordance with the results from batch flotation tests as shown in Figure 2. This suggests that most of the serpentine slime may report to the concentrate by becoming trapped in or coated on these hydrophobic flocs of ultra-fine metal minerals.

Effects on entrainment of serpentine by combined collectors – microphotographic measurements Figures 4a and 4b show photographs of suspensions of the concentrates from batch flotation tests. These images show some loose and fluffy hydrophobic flocs (dark colour) in suspension, coupled with some serpentine slimes (light grey colour) either by entrapment in or coating on these hydrophobic flocs. This indicates that both coating and entrapment may contribute to the entrainment of serpentine in concentrates. However, since the recoveries and grades of copper and nickel in these concentrates, as well as the MgO content, are only slightly affected by high-intensity conditioning (an increase in the stirring speed from 800 to 1600 r/min in an attempt to remove the slime coating), it is likely that entrapment makes a greater contribution to serpentine entrainment than coating. It is noteworthy that this mechanism of slimes coating on the relatively coarse (>37 µm) sulphide particle surfaces might be more favourable for serpentine entrainment than the other due to very little hydrophobic flocs being formed by strong collectors. These results agree well with those illustrated in Figures 2 and 3.

Possible mechanisms of entrapment and entrainment of serpentine in concentrates

Figure 3 compares the settling rates of a head feed and a decollector concentrate before and after treatment by the combined collectors (BX plus BA) at the same dosages that were used in the batch flotation tests. The data indicates that in the first 8 minutes, the settling rates of both the head feed and concentrate were significantly faster when treated with combined collectors. Furthermore, the settling rate of this concentrate (treated with the collectors) is much faster than the others, suggesting that the combined collectors increase the sizes of these fine metal mineral aggregates by forming

Figure 5 compares the recoveries of Cu, Ni, and Pt+Pd at different size fractions in the concentrate from batch flotation tests based on the flow sheet shown in Figure 1. High recoveries (>80% of Cu, Ni, and Pt+Pd) were obtained in the -37 µm size fraction, particularly in the -20 µm fraction, but the MgO content of the concentrate remained almost 19%.

These results suggest that the ultra-fine copper and nickel sulphides in the ore samples can be effectively flocculated by using the strong combined collectors (BX plus BA) due to the formation of fluffy hydrophobic flocs that attach to the

Figure 2 – Effect of combined collector dosage on the recovery and grade of Ni in batch flotation tests

Figure 3 – Settling rates of concentrate and head feed before and after treatment with the combined collectors (BX plus BA)

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Effects of combined collectors on entrainment of serpentine – settling rate measurements

Effects on entrainment of serpentines by hydrophobic flocs of ultra-fine copper-nickel sulphides during flotation ratio) can effectively flocculate the ultra-fine metal sulphide minerals, forming a certain amount of loose and fluffy hydrophobic flocs that attach to the bubbles. However, these flocs might correspondingly provide many potential spaces for serpentine entrainment into the concentrate by either entrapment in or coating on the flocs, resulting in high MgO contents (up to 19%) in the concentrate. The entrapment of serpentine slimes in the hydrophobic flocs may play an important role in contaminating the concentrate at a size range of less than 37 µm. Entrapped serpentine slimes will be more difficult to remove that slimes coated onto sulphide particles, which can be removed by high-intensity conditioning or using efficient depressant/flocculants.

Acknowledgements

Figure 4 – Microphotographs of concentrate suspension from batch flotation tests

MT acknowledges the Wenbin Zhang’s project on ultra-fine copper-nickel sulphides containing platinum metal groups in Yunnan, which had been completed. Financial support for this project was provided by the Natural Sciences Council of China. The authors also appreciate support from the Scholarship of Yunnan Science.

References BOZKURT, V., XU, Z., and FINCH, J.A. 1998. Pentlandite/pyrrhotiite interaction and xanthate adsorption. International Journal of Mineral Processing, vol.

52. pp. 203–214.

BULATOVIC, S.M. 1999. Use of organic polymers in the flotation of polymetallic ores: a review. Mineral Engineering, vol. 12. pp. 341–354.

BULATOVIC, S.M. 2003. Evaluation of alternative reagent schemes for the flotation of platinum group minerals from various ores. Minerals Engineering, vol. 16. pp. 931–939.

CHEN, G., GRANO, S., SOBIERAJ, S., and RALSON, J. 1999. The effect of high intensity conditioning on the flotation of a nickel ore, paper 2: mechanisms. Mineral Engineering, vol. 12. pp. 1359–1373.

Figure 5 – Distribution of Ni, Cu, and Pt+Pd in different size fractions in the concentrate from batch flotation tests

EDWARDS, C.R., KIPKLE, W.B., and AGAR, G.E. 1980. The effect of slime coatings of the serpentine minerals, chrysotile and lizardite, on pentlandite flotation. International of Journal of Mineral Processing, vol. 7. pp. 33–42.

FENG, B., FENG, Q., LU, Y., and LV, P., 2012. The effect of conditioning methods and chain length of xanthate on the flotation of a nickel ore. Minerals Engineering, vol. 39. pp 48–50.

bubbles and report to the concentrate. However, the high serpentine content of this concentrate makes these results unsatisfactory. These collectors have little influence on the floatability of serpentine minerals, according to previous research by Lu et al. (2011). Based on the results from microphotographs and settling rate measurements, the possible mechanisms of serpentine entrainment may relate to both entrapment in and coating on these fluffy hydrophobic flocs of ultra-fine copper and nickel sulphides. An attempt to remove the entrapped and entrained serpentine by using intensified conditioning ( 1600 r/min for 30 minutes) before batch flotation resulted in a very slightly decrease of MgO grade in the concentrate. That indicates that serpentine entrapment in metal sulphide flocs may play much more important role than coating on the particles. Those results agree well with previous research (Wellhame et al., 1992; Senior et al., 1994; Senior and Thomas, 2005). Certainly, serpentine recovery by water recovery should not be ignored, but is probably less important than other mechanisms after using CMC and KN as depressants.

Conclusions Based on the results of batch flotation and microflotation tests, the combined collectors (BX plus BA at a 1:2 weight

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GLEMBOTSKII, A.A. 1958. The combined action of collectors during flotation.

Tsvetnye Metally, vol. 4. pp. 6–14.

LOTTER, N.O. and BRANDSHAW, D.J. 2010. The formulation and use of mixed collectors in sulfide flotation. Minerals Engineering, vol. 23. pp. 945–951.

LU, Y., ZHANG, M., FENG, Q., LONG, T. L., OU, L., and ZHANG G. 2011. Effect of sodium hexametaphosphate on separation of serpentine from pyrite.

Transactions of Nonferrous Metals Society of China, vol. 21. pp. 208–213.

MALYSIAK, V., O’CONNOR, C.T., and RALSTON, J. 2002. Pentlandite-feldspar interaction and its effect on separation by flotation. International Journal of Mineral Processing, vol. 66. pp. 89–106.

SENIOR, G.D., SHANNON, L.K., and TRAHAR, W.J. 1994. The flotation of pentlandite from pyrrhotite with particular reference to the effects of particle size.

International Journal of Mineral Processing, vol. 42. pp. 169–190.

SENIOR, G.D. and THOMAS, S.A. 2005. Development and implementation of a new flowsheet for the flotation of a low grade nickel ore. International Journal of Mineral Processing, vol. 78. pp. 49–61.

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The Journal of The Southern African Institute of Mining and Metallurgy

http://dx.doi.org/10.17159/2411-9717/2015/v115n8a5

Smelting of calcined basic nickel carbonate concentrate in a 200 kW DC arc furnace

by M. Abdellatif*

Calcined basic nickel carbonate (BNC) concentrate was smelted in a pilotscale DC arc furnace to produce a nickel metal. The furnace was continuously operated for 12 days (24 hour/day), during which twelve different smelting conditions were investigated, with the major variables being reductant type and feed rate, flux composition and addition, and BNC feed rate.

The 200 kW DC arc furnace was operated at power levels between 110 and 165 kW and at a total feed rate of 78 to 96 kg/h, resulting in an average slag and metal tapping temperature of about 1650°C. A total of 7.2 t of BNC were smelted, producing about 5.44 t of nickel metal and 2.94 t of slag. Nickel recoveries of 96.4% and higher were achieved, and the slag nickel content was as low as 0.1%. The major impurities in the metal were iron (mostly from oxygen lancing) and carbon. The calculated feed carry-over was less than 0.85% and the graphite electrode consumption was between 2.8–3.3 kg/MWh. Keywords nickel smelting, DC arc furnace, calcined basic nickel carbonate, BNC.

Introduction One of the most distinctive advantages of DC open-arc smelting is the ability to process fine raw materials without any major issues with regard to the operability of the furnace, metal recovery, and metal quality. This has been demonstrated in Mintek’s DC arc pilot plant facilities over a period of about three decades (Curr et al., 1983). The materials that have been processed include laterite (Kotze, 2002), chromite fines (Curr et al., 1983), manganese fines (Lagendijk et al., 2010), PGM concentrates (Shaw et al., 2013), electric arc furnace (EAF) dust (Denton et al., 2005; Abdellatif, 2002a) and stainless steel (SS) dust (Denton et al., 2005; Abdellatif, 2002a; Goff and Denton, 2004), and petroleum fly ash (Abdellatif, 2002b). The top particle size of such materials can vary from a few millimetres (e.g. laterite) to micrometres (EAF and SS dust). Dusting, whether due to physical carryover of feed or to the arc side reactions, has been proven to be a non-issue. As such, the applications of DC open-arc furnaces have the potential to be extended to other raw materials such as particulate nickel oxide, which has a particle size much less than 100 μm.

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DC arc pilot plant The 200 kW DC arc pilot plant consists of a DC power supply, refractory-lined furnace, and an off-gas treatment system (Figure 1). The furnace comprises a refractory-lined cylindrical steel shell, a domed base, and a conical roof.

The furnace shell has an unlined internal diameter of 980 mm and is provided with water-spray cooling. The conical roof contains

* Mintek, Randburg. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Jan. 2014 and revised paper received May 2014.

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Synopsis

Nickel oxide is typically produced from laterite ore by sulphuric acid leaching, purification, and precipitation as basic nickel carbonate (BNC) (Rhamdhani, Jak, and Hayes, 2008) or as mixed hydroxide precipitate (MHP) (Mackenzie, Virnig, and Feather, 2006), and finally calcining at moderate temperatures (800–1200°C). The NiO product is very fine, with an average particle size in the range of 10–20 μm and a NiO content of more than 99%.

Nickel oxide can be reduced to nickel metal by electrolysis (Moskalyk and Alfantazi, 2002), hydrogen reduction (Agrawal et al., 2006), or by the carbonyl process (Terekhof and Emmanuel, 2012). An alternative approach would be to smelt the NiO in a DC open-arc furnace in the presence of slag fluxes and a carbonaceous reducing agent. This alternative might offer faster reaction kinetics, and thus a smaller processing unit, as compared with hydrogen or CO reduction. In addition, it provides almost prompt metal separation from the slag, which allows the metal to be cast as required. The DC option may offer lower capital costs in comparison to the electrolytic process, as well as minimizing environmental pollution. This approach was investigated using Mintek’s 200 kW DC arc facility. The results of the investigation are summarized in this paper.

Smelting of calcined basic nickel carbonate concentrate in a 200 kW DC arc furnace Experimental procedure The initial warm-up period involved melting a 75 kg nickel heel, using cathode-grade nickel billets supplied by Insimbi Alloy Supplies, and charging two batches of raw materials into the furnace. The various feed components were then fed at the target values, and the power set-point adjusted to match the total feed rate and energy loss set-points. Feeding continued for between 1.5 and 3.0 hours, depending on the batch mass and the total feed rate. The metal and slag were then tapped into refractory-lined steel ladles by oxygen lancing. After initial cooling, the ladles were weighed and removed for further cooling and subsequent metal/slag separation, weighing, and storage.

The metal and slag temperatures were measured during tapping using an infrared pyrometer. Spoon samples were taken from both the metal and the slag stream. The slag samples were prepared for chemical analysis by crushing to about -15 mm, visually separating the metallic inclusions, and pulverizing to -75 μm. The metal samples were obtained by taking drill shavings using either a titanium or stainless steel drill bit.

The entrained solids in the furnace off-gas were captured in a scrubber unit using 1% NaOH solution. The scrubber slurry was removed every 24 hours, sampled, weighed, and analysed. Very little baghouse dust was recovered throughout the test work. When available, the dust was weighed, sampled, and analysed.

Samples of the various products (and the feed components) were chemically analysed using an ICP-OES method. Carbon and phosphorous were determined using a Leco technique.

The flux was prepared by mixing weighed quantities of the components using a Jones mixer. All operational data was monitored and recorded every two seconds on a PLC datalogging system.

The entire test work was carried out as one continuous campaign which lasted for 12 days (24 hours a day), and comprised 12 smelting conditions (Table I), in which the main variables were flux composition and addition rate, type and feed rate of reducing agent, and the total feed rate. The smelting conditions were guided by thermodynamic calculations carried out using the Pyrosim modelling software (Jones, 1987).

Figure 1 – Schematic layout of the 200 kW DC arc facility

two ports for feeding and a central port for a 100 mm graphite electrode (cathode). The roof is cooled by means of pressurized water panels.

The furnace shell was lined with magnesia bricks, giving an internal diameter of about 670 mm. The base and the roof were lined with alumina castable.

The return electrode (anode) consists of several steel rods built into the hearth refractories. These rods are connected at their lower ends to a steel plate attached to the hearth dome, which is connected to the anode busbars.

The feed system comprised three feeders: the primary BNC feeder, the flux feeder, and a reductant feeder. Each hopper was equipped with load cells and a variable speed vibratory feeder in order to enable a controlled feed rate to be delivered to the furnace. The feed rate control was linked to the Delta V control and data acquisition system.

The power supply consisted of an 11 kV vacuum breaker, two isolators, two contactors, two transformers, and two 5 kA DC thyristor rectifiers. The gas extraction system comprised trombone coolers, a scrubber, a reverse-pulse bag filter, a fan, and a stack.

Table I

Test conditions Taps

1-8 9-10 11-21 22-27 28-31 32-34 36-46 47-52 53-59 60-64 65-69 70-73

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692

Condition

1 2 3 4 5 6 7 8 9 10 11 12

Reductant

Flux

BNC feed rate

Power set-point

Feed-on time

BNC fed

Type

% of BNC

Type

% of BNC

kg/h

kW

h

kg

Sascarb Sascarb Anthracite Sascarb Sascarb Sascarb Sascarb Sascarb Sascarb Sascarb Met. coke Met. coke

11.2 11.2 15.0 12.0 13.0 14.0 14.0 16.0 14.0 14.0 16.8 16.8

1 1 1 1 1 1 2 2 2 2 2 2

40 40 40 40 40 40 40 40 30 20 40 40

50 50 50 50 50 50 50 50 50 50 50 60

80-130 125 125-150 140-150 140 140-150 150 155 155 140-150 145-150 145-165

3 2 2 2 2 2 2 2 2 2 2 1.5

842.4 177.7 1013.8 592.7 386.4 285.9 1127.6 612.9 706.7 504.7 497.6 364.1

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Smelting of calcined basic nickel carbonate concentrate in a 200 kW DC arc furnace Results

The raw materials employed in the test work consisted of eight different components. The nickel source was calcined basic nickel carbonate (BNC) concentrate. The average analysis of the BNC is shown in Table II. The particle size averaged about 10 μm, with a top size of about 40 μm.

Three different reducing agents were utilized, namely petroleum coke (referred to as Sascarb), anthracite coke, and metallurgical coke (Table III). Sascarb is a relatively pure carbon source (>99% C) and has a narrow particle size range of 3–5 mm. The Sascarb was acquired from Sasol. Anthracite coke has a relatively low fixed carbon content (79.7% C), with a particle size of -10 mm. The metallurgical coke analysed at about 84% fixed carbon and had a particle size of less than 10 mm.

The flux used in Conditions 1-6 (Table I) consisted of lime, magnesia, and silica (Table II), where the target basicity ratio was kept at 1.2–1.0 (defined as (CaO + MgO)/SiO2). The particle size of both lime and magnesia was -5 mm, while that of silica was between 2 mm and 10 mm. In Conditions 712, alumina was added to the flux mixture with a target A12O3 content in the slag of 20%, while keeping the basicity ratio at 1.2–1.0 (defined as (CaO + MgO)/ (SiO2 + Al2O3)).

The alumina used had a particle size of -10 mm.

Table II

Average chemical analyses of the BNC and fluxing agents, mass % Component

BNC

Lime

Magnesia

MgO 0.031 2.056 92.164 CaO 0.001 90.10 2.014 Al2O3 0.038 0.395 0.722 SiO2 NA 1.524 0.720 Fe2O3 0.038 0.357 1.284 MnO 0.051 1.023 0.410 NiO 97.439 0.004 0.011 S 0.328 0.040 0.010 P 0.022 0.013 0.040 Co, ppm 195 NA 10 LOI < 0.10 4.321 2.400 * Includes moisture content. NA: not analysed.

Silica

Alumina

0.049 0.073 0.948 98.215 0.598 <0.001 0.008 0.010 0.010 10 NA

NA 5.350 94.20 0.100 0.10 0.064 NA NA NA NA NA

Table III

Average chemical analyses of the reductants, mass % Component

Sascarb

Anthracite

MgO < 0.083 0.257 CaO <0.07 0.322 Al2O3 0.321 2.229 SiO2 0.285 6.064 Fe3O3 <0.072 0.786 NiO, ppm <10 <57 CoO 0.656 <0.002 LOI* <0.10 8.892 C 99.340 79.770 P 0.003 0.016 S 0.112 0.180 * Includes moisture content. NA: not analysed.

Met. coke 0.182 0.280 2.456 6.161 0.357 <57 <0.064 5.575 84.20 0.011 0.501

The Journal of The Southern African Institute of Mining and Metallurgy

Slag analyses The weighted average slag analysis for the various conditions is presented in Table IV (see also Figure 2 for nickel oxide range of analysis). With the flux-1 recipe, the nickel content in the slag remained relatively high for the first six conditions, ranging from about 0.60% to 10.1% NiO, although in certain taps it dropped to below 0.5%. (Condition 2 is excluded from the analysis as it consisted of only two taps). This may be attributed to several factors. ➤ Metallic inclusions in the samples. This was evident from the analysis of cleaner samples taken from the ladle, which had a significantly lower nickel content.

(Ni analyses of certain spoon sample taken during tapping were 50% higher than analyses of the corresponding clean samples taken from the ladle.)

Incomplete metal-slag separation in the furnace was attributed largely to the viscous and sticky slag ➤ The average furnace tapping temperatures were generally lower than those achieved in subsequent conditions ➤ Possibly insufficient reductant.

Replacing Sascarb with anthracite while keeping the fixed carbon addition constant (Condition 3) did not have a major impact on the nickel content in the slag, as compared to Condition 1. However, if the results of taps 1-3 (warm-up period) are excluded from Condition 1, the use of anthracite appears to decrease the slag nickel content in certain taps to about 1.0%, compared to a low of 4.3% realized in Condition 1. In Conditions 4-6, the Sascarb addition was gradually increased from 12% to 14% of the BNC mass, compared to 11.2% in Condition 1. This resulted in a significant decrease in nickel oxide content in the slag, to an average of about 4% in Conditions 5 and 6. The minimum values achieved in certain taps during these three conditions were not significantly different, ranging from 1.22 to 1.86% NiO. Condition 7 represents the start of the flux-2 recipe (addition of alumina). The Sascarb addition was kept at 14% (the same as for Condition 6). The resulting slag had the lowest average NiO content (about 0.88%) thus far. Further increase in the reductant addition to 16% (Condition 8) yielded a slag containing less than 0.5% NiO in certain taps.

With lower flux addition (30% of the BNC – Condition 9), nickel oxide in the slag averaged 0.56%, and reached a level of 0.13% in more than one tap. A further decrease in the flux addition in Condition 10 appears to have resulted in an increase in the NiO content (1.38% average, with the lowest being 0.41%).

Metallurgical coke was employed as the reducing agent in Condition 11 (16.8% addition to give a similar fixed carbon addition as in Condition 7). Simultaneously, the flux-2 addition was increased to 40% of the BNC mass. The slag produced contained between 0.2 and 2.1% NiO. These results suggest that the metallurgical coke was somewhat more effective than Sascarb in reducing the nickel oxide. The feed recipe for Condition 12 was similar to that for Condition 11, but the feed rates were increased by 20%. The nickel oxide content in the slag did not change significantly from that obtained in Condition 11, averaging about 0.6%, with a minimum value of 0.18%.

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Raw materials

Smelting of calcined basic nickel carbonate concentrate in a 200 kW DC arc furnace Table IV

Weighted average chemical analyses of the slag, mass % Condition 1 2 3 4 5 6 7 8 9 10 11 12 Dig-out

Slag, kg

NiO

Al2O3

MgO

SiO2

CaO

FeO

S

Total

376.0 48.9 494.0 264.5 152.4 96.7 342.7 263.2 252.0 120.3 255.3 140.7 129.6

9.03 0.61 10.07 7.92 3.98 3.90 0.88 1.25 0.56 1.38 0.94 0.63 5.36

1.30 0.89 1.55 2.47 1.61 1.74 18.11 19.76 19.97 20.32 29.69 21.00 20.21

14.79 17.77 16.85 16.96 16.90 19.71 17.11 32.71 16.38 15.75 14.89 16.65 19.57

49.97 51.24 41.08 40.82 45.11 43.59 28.14 22.46 22.93 23.23 20.12 45.02 15.86

29.23 31.14 27.51 29.32 30.80 32.87 37.81 39.90 39.86 40.88 31.82 32.05 36.26

4.18 0.86 3.27 2.61 2.08 0.90 0.35 0.76 0.57 0.68 1.02 0.30 1.03

0.03 0.04 0.04 0.01 0.01 0.02 0.03 0.08 0.12 0.22 0.11 0.51 0.60

105.58 102.55 100.37 99.98 100.49 102.73 102.43 116.92 100.39 102.46 98.59 116.16 98.89

Metal analyses

Figure 2 – NiO analyses of the slag per condition

The slag from certain taps had a relatively high FeO content of up to 13.8%. This was most likely due to excessive lancing, which was required on a few occasions. The average total slag composition exceeds 100% in most of the conditions (Table IV). This might be related to metallic inclusions, as well as analytical errors.

The weighted average chemical analyses of the metal are shown in Table V. The nickel produced in Conditions 1-6 (flux-1 recipe) analysed at between about 0.9% and 4.5% impurities, the main impurity element being iron. This is believed to be due to oxygen lancing, in addition to the iron present in the feed materials. The second major impurity is cobalt at about 0.10–0.34%. The carbon content in the metal varied between 0.01% and 0.42% except for Condition 6, where it increased to about 1.0% as a result of the increased carbon addition.

Condition 7 produced a cleaner metal in terms of iron content, as lancing was very moderate throughout most of the taps. However, the carbon content averaged about 2.0%.

Further increase in the reductant addition (Condition 8) yielded lower metallic impurity levels, totalling about 3.3%.

In this condition the iron content dropped to about 0.2%. In Condition 9 (30% flux-2 and 14% Sascarb), similar quality metal to that of Condition 8 was produced, with only a marginal increase in both carbon and sulphur contents.

Carbon and iron were the major impurity elements in the nickel produced in Condition 10, averaging about 1.8% and 1.6% respectively. Similar quality metal was produced in Condition 11, where metallurgical coke was employed, except

Table V

Chemical analyses of nickel metal, mass % Condition

Metal, kg

Si

Ca

Mn

Fe

Co

C

S

Ni

1 2 3 4 5 6 7 8 9 10 11 12 Dig-out

351.0 271.1 679.4 388.9 318.9 329.0 899.0 419.6 564.1 547.5 234.7 195.7 243.0

0.17 0.27 0.23 0.05 0.05 0.05 0.06 0.05 0.05 0.05 0.09 0.05 0.69

0.32 0.21 0.23 0.07 0.34 0.15 0.27 0.21 0.15 0.26 0.34 0.072 0.27

0.05 0.05 0.05 0.05 0.05 0.05 0.13 0.18 0.16 0.11 0.13 0.09 0.44

3.63 1.72 0.88 0.93 1.06 4.54 0.53 0.21 0.24 1.65 0.96 1.39 2.23

0.15 0.09 0.34 0.34 0.34 0.28 0.11 0.31 0.33 0.30 0.32 0.34 0.24

0.10 0.42 0.15 0.01 0.06 0.99 2.01 1.92 2.14 1.78 2.34 1.41 1.49

0.067 0.20 0.03 0.018 0.018 0.049 0.049 0.041 0.061 0.066 0.102 0.028 0.077

95.22 96.83 97.62 98.08 97.67 93.45 96.42 96.68 96.47 95.35 95.29 96.20 94.34

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Smelting of calcined basic nickel carbonate concentrate in a 200 kW DC arc furnace

Fume analyses The major component in the fume product is nickel oxide (about 48%, Table VI), making it suitable for recycling to either the furnace or the calciner, depending on the sulphur content and other impurities. Semi-quantitative XRD analysis suggested that the crystalline phases are 32% Ni, 51% NiO, 12% Ni3S2, and 5% CaCO3.

Scrubber products The normalized sludge analysis appears to be similar to that of the fume, particularly with regards to nickel oxide and sulphur contents (Table VII). Therefore, it might be possible to recycle the sludge to the calciner in order to remove the moisture and to minimize the sulphur content before feeding it to the furnace.

Several samples were taken from the effluent stream. The nickel content was well below 2 ppm in all of these samples, and therefore the effluent stream required only neutralizing before disposal.

Feed carry-over The proportion of the feed materials that escaped the furnace with the off-gas was calculated via two approaches. The first assumed that all the fume and scrubber sludge (dry basis) were unreacted materials and therefore represented the total losses. This approach results in a feed carry-over of about 0.85%. The second approach was based on the calcium content in the fume and scrubber sludge relative to that in the feed materials. Specifically, the feed carry-over was calculated from the following relationship: Carry-over = Mass% Ca in fume × Fume masses × 100% /(Mass% Ca in feed × Total mass fed) This approach resulted in a feed carry-over of about 0.5%, which is significantly less than that of the first

Table VI

Chemical analyses of the fume, mass % MgO 7.67 Al2O3 1.68 SiO2 6.10 CaO 5.71 MnO 1.04 P 0.01 * Balance is mostly moisture

Fe2O3 NiO CoO C SO3 Total*

0.97 48.17 0.46 0.47 12.73 85.55

approach. Nevertheless, both methods suggest that the feed losses to the off-gas were very low, particularly given the scale of operation and the particle size of the feed components.

Overall mass balance The overall mass balance for the entire test campaign shows an accountability of about 103%, or about 308 kg of excess products, as compared to the masses fed (Table VIII). This is believed to be related to the CO/CO2 mass ratio of 1.22/1.0 as predicted by Pyrosim (Jones, 1987), oxygen lancing, which introduced more than 368 kg of iron into the furnace products, and refractory erosion.

For the first flux recipe, the overall accountability was 93.7%, representing a deficit of about 333 kg. This is mostly slag and metal build-up in the furnace that could not be tapped at the end of this period. Conversely, the flux-2 recipe exhibited an overall accountability of about 111%, or a surplus of about 640 kg.

Nickel mass balance For the flux-1 recipe, the nickel accountability was about 90.3%, compared to about 101.0% for the flux-2 recipe.

Inclusion of the metal digout in the flux-2 recipe period is believed to be the major reason for this. Based on mineralogical examination of three slag samples, metallic nickel in the slag ranged from about 2.1% to 5.4% (mass basis), averaging about 4.23% in the samples tested. This is equivalent to about 125 kg of metal trapped in the slag.

Taking this into account, the overall nickel accountability improves to over 98.2%.

Nickel recoveries averaged about 96.4% and 99.5% for the flux-1 test work and 88.8 and 99.7%, for flux-2, for metal and slag masses and analyses respectively.

Energy balance and electrode consumption Based on the measured heat losses and the total energy input, the furnace thermal efficiency averaged about 36.5% in Conditions 1–6. It dropped to about 29.5% in Conditions 7–12, possibly due to refractory erosion. It should be noted that the furnace power intensity averaged about 300 kW per square metre of furnace hearth area during the first six conditions. It then increased to 390 kW/m2 for the rest of the test work, except for Condition 12 where it averaged about 450 kW/m2.

Table VIII

Overall mass balance for the two flux recipes In, kg Flux 1

Table VII

Chemical analyses of the sludge, mass % MgO 4.32 Al2O3 1.01 SiO2 3.29 CaO 4.18 MnO 0.72 NiO 28.90 *Balance is mostly moisture

CoO C SO3 Na Fe2O3 Total*

0.18 1.38 10.20 0.40 0.42 53.23

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BNC Reductant Silica Magnesia Lime Alumina Ni heel Total Accountability Surplus (deficit)

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Out, kg Flux 2

3298.9 3909.5 436.5 576.7 668.8 295.7 221.4 215.9 590.3 605.1 0.0 277.3 75.0 0.0 5290.8 5880.3 93.7% 110.9% (333.4kg) 639.0kg

Slag Metal Fume Off-gas Sludge

Flux 1

Flux 2

1432.2 2356.4 30.5 1138.4 0.0

1503.8 3085.5 64.9 1811.7 53.4

4957.5

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for the iron and carbon contents, which averaged about 1.0% and 2.3% respectively. With a higher smelting rate during Condition 12, iron content increased to about 1.4%, while the level of carbon decreased to 1.4% compared to 2.3% Condition 11.

Smelting of calcined basic nickel carbonate concentrate in a 200 kW DC arc furnace The electrode consumption was 3.3 kg/MWh in the first six conditions. Longer arc length, and thus lower current density, and more steady operation during the flux-2 test work contributed to the almost 38% drop in electrode consumption (2.8 kg/MWh), as compared to the flux-1 recipe.

Discussion Although the nickel recovery during flux-2 test work was very high at fixed carbon additions of 14% of BNC or higher, it is believed that further optimization can be achieved. A larger and better sealed furnace may reduce air ingress, and thus carbon losses through oxidation. The carbon addition could be adjusted to control the carbon content of the metal and to increase the slag sulphur capacity, in addition to controlling the reduction of nickel oxide. These are somewhat contradictory objectives, but a balance may need to be found.

Proper selection of the reducing agent and the flux components can help reduce the sulphur content in the metal (Pashkeev et al., 2011; Shankar, 2006; Ren, Hu, and Chou, 2013; Wang et al., 2009), and thus its refining requirements, if any.

Smelting of the BNC in the presence of a CaO-MgO-SiO2 slag (flux-1 test work) was somewhat difficult and resulted in slags with relatively high nickel contents. However, increased fixed carbon additions tended to increase nickel recovery, as is shown by the drop in the average nickel content in the slag. Nickel losses to the slag are believed to be affected by metallic inclusions, as the slag tended to be sticky. The formation of high-melting compounds (carbides) could have affected the flowing characteristics of the slag and hindered the settling of the metal.

With flux-2 recipe there did not appear to be a clear relationship between the average nickel content of the slag and the fixed carbon addition within the range studied. The lowest carbon addition during this period (about 13.6%) was sufficient to lower the nickel levels to below 1.0%, on

average. Any extra carbon added is believed to have resulted in the reduction of other metal oxides, and more importantly in the accumulation of reductant in the furnace. In addition, elevated reductant additions tended to increase the carbon content in the product metal (Figure 3).

In addition to enhancing nickel extraction, higher carbon additions were aimed at lowering the oxygen potential in order to increase the sulphur capacity of the slag, and therefore decrease the sulphur content of the metal. The results, however, did not show a strong relationship between the average sulphur content in the metal and fixed carbon addition. As indicated previously, the slag/metal ratio, slag depth, slag composition, and operating temperature were not constant throughout the test work, which would account for the apparent lack of such correlation. For example, when the results of Conditions 7, 9, and 10 are examined separately (flux additions of 40, 30, and 20% respectively), it could be concluded that increased flux addition tended to lower the sulphur content of the metal. At the same time, the mass ratio (% Sslag/% Smetal) tended to drop with increased flux addition, mostly due to dilution effects (Figure 4). Overall, and based on the sulphur mass balance, about 19.8% and 20.4% of the sulphur in the feed reported to the slag and metal respectively, with the balance being in the fume, sludge, and the scrubber effluent. The sulphur accountability is about 104%.

Conclusions Particulate BNC concentrate was successfully smelted in a 200 kW DC arc furnace to produce a metal containing 96% Ni and higher. Two slag recipes were employed. The first consisted of 40% CaO, 15% MgO, and 45% SiO2. Nickel recoveries were acceptable and averaged about 96.4%, with the slag containing 8.4% NiO on average. The recovery of the metal was improved with increased reductant additions and higher operating temperatures.

Figure 3 â&#x20AC;&#x201C; Fixed carbon addition vs carbon content in the metal (Conditions 1-12)

â&#x2013;˛

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Smelting of calcined basic nickel carbonate concentrate in a 200 kW DC arc furnace

Figure 4 – Influence of flux addition on sulphur behaviour, Conditions 7, 9, and10

Nickel recovery to the metal phase increased to 99.7% when the second slag recipe was used. The produced slag (40% CaO, 15% MgO, 25% SiO2, and 20% Al2O3) contained less than 1% NiO on average. Nickel recovery did not vary significantly, regardless of the type of reductant, total feed rate, or the amount of flux used.

Sulphur retention in the slag appears to have been influenced by the slag composition and, to a lesser extent, by the reductant addition. Carbon content in the metal tended to increase with higher reductant additions, regardless of the slag composition or the nature of the reductant.

Despite the very fine particle size of the BNC, feed carryover was well below 0.9% throughout the test work. The fume contained about 40% NiO, and can be easily recycled to the calcining step in order to minimize the sulphur content and to drive off the moisture.

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ABDELLATIF, M. 2002b. Recovery of vanadium and nickel from petroleum flyash.

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JONES, R.T. 1987. Computer simulation of pyrometallurgical processes. APCOM ’87. Proceedings of the Twentieth International Symposium on the Application of Computers and Mathematics in the Minerals Industries, Johannesburg, South Africa. Vol. 2: Metallurgy. South African Institute of Mining and Metallurgy, Johannesburg. pp. 265–279.

KOTZE, I.J. 2002. Pilot plant production of ferronickel from nickel oxide ores and dust in a DC arc furnace. Pyrometallurgy 02, Mount Nelson Hotel, Cape Town, South Africa, 11-12 March 2002.

LAGENDIJK, H., XAKALASHE, B.S., LIGEGE, T., NTIKANG, P., and BISAKA, K. 2010.

Comparing manganese ferroalloy smelting in pilot scale AC and DC submerged arc furnaces. INFACON XII. Proceedings of the 12th International Ferro-Alloys Congress, Helsinki, Finland, 6–9 June 2010. pp. 497–507.

MACKENZIE, M., VIRNIG, M., and FEATHER, A. 2006. The recovery of nickel from high-pressure acid leach solutions using mixed hydroxide product-LIX(R) 84-INS technology. Minerals Engineering, vol.19, no. 12. pp. 1220–1233.

MOSKALYK R.R. and ALFANTAZI, A.M. 2002. Nickel laterite processing and electrowinning practice. Minerals Engineering, vol. 15. pp. 593–605.

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RHAMDHANI, M.A., JAK, E., and HAYES, P.C. 2008. Basic nickel carbonate; Part I. Microstructure and phase changes during oxidation and reduction processes. Metallurgical and Materials Transactions B, vol. 39B. pp. 218–228.

SHAW, A., DE VILLIERS, L.P.VS., HUNDERMARK, R.J., NDLOVU, J., NELSON, L.R., PIETERSE, B., SULLIVAN, R., VOERMANN, N., WALKER, C.. STOBER, F., and MCKENZIE, A.D. 2013. Challenges and solutions in PGM furnace operation: high matte temperature and copper cooler erosion; Journal of the Southern African Institute of Mining and Metallurgy, vol. 113, no. 3. pp. 251–26

CURR, T.R., BARCZA, N.A., MASKE, K.U., and MOONEY, J.F. 1983. The design and operation of transferred-arc plasma systems for pyrometallurgical application. 6th International Symposium on Plasma Chemistry (ISPC-6), Montreal, Canada, July 1983.

TEREKHOV, D.S. and EMMANUEL, N.V. 2012. Direct recovery of nickel and iron from laterite ores using the carbonyl process. Proceedings of Processing of Nickel Ores and Concentrates’12, Cape Town, South Africa, November 2012.

DENTON, G.M., BARCZA, N.A., SCOTT, P.D., and FULTON, T. 2005. EAF dust processing. John Floyd International Symposium on Sustainable Developments in Metal Processing, Melbourne, Australia, 3–6 July 2005.

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AGRAWAL, A., KUMAR, V., PANDEY, B.D., and SAHU, K.K. 2006. A comprehensive review on the hydrometallurgical process for the production of nickel and copper powders by hydrogen reduction. India Materials Research Bulletin, vol. 41. pp. 879–892.

PRE-CONFERENCE WORKSHOPS Pre-conference workshops will be organised in conjunction with World Gold 2015: Gold Processing Workshop

2015 Building a Resilient Gold Mining Industry For further details contact: SAIMM, Conference Co-ordinator, Camielah Jardine Tel: 27 (11) 834-1273/7 Facsimile 27 (11) 838-5923 E-mail: camielah@saimm.co.za Website: www.saimm.co.za

28 September 2015 · Workshop 29 September–1 October 2015 · Conference 2 October 2015 · Technical Visits Misty Hills, Gauteng, South Africa Incorporating geology, metallurgy and mining

TOPICS World Gold 2015 will reflect on these key issues with greater focus on improved efficiency and the latest technology in: ➺ geological mapping ➺ gold mining ➺ mineral processing, and ➺ extraction and refining ➺ human resources ➺ financial resources ➺ computer assisted exploration targeting in gold exploration ➺ application of partial extraction methods in gold exploration ➺ geochemical and/or mineralogical haloes as indicator for gold targets ➺ ‘Black Smokers’ as exploration targets for gold ➺ are there still mega gold deposits waiting to be discovered? ➺ Brownfields gold exploration, as success story.

Thus, we are focusing on more environmentally friendly, resource efficient and energy efficient mining and recovery methods.

The Southern African Institute of Mining and Metallurgy (SAIMM), the Canadian Institute of Mining, Metallurgy and Petroleum (CIM) and the Australasian Institute of Mining and Metallurgy (AusIMM) will jointly convene a World Gold Conference every two years. In 2015 it will be held in Johannesburg, South Africa and hosted under the auspices of the SAIMM. Some important aspects of the current mining environment will provide opportunities and threats for the industry in the foreseeable future, which include: ➺ Gold price volatility ➺ Minimal exploration success for the last 10 years and little immediate prospect for revolutionary success is leading to revisiting known ‘old’ (sub-marginal) deposits ➺ Lower precious metal content Keynote Speakers: L. Lorenzen, Mintrex ➺ Increasing refractoriness M. Reuter, Outotec ➺ More energy efficient mining and processing M. Marcombe, AngloGold ➺ Decreasing equity risk for juniors and midAshanti tiers ➺ Maximising long-term optionality

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http://dx.doi.org/10.17159/2411-9717/2015/v115n8a6

Comparison of physical properties of oxidative sintered pellets produced with UG2 or metallurgicalgrade South African chromite: a case study

by R.I. Glastonbury*, J.P. Beukes*, P.G van Zyl*, L.N. Sadiki†, A. Jordaan*, Q.P. Campbell*, H.M. Stewart‡, and N.F. Dawson‡

The physical properties of oxidative sintered pellets produced from typical South African UG2 ore are compared with the physical properties of pellets produced with conventional South African metallurgical-grade chromite ore (from the Lower Group 6 or the Middle Group 1 and 2 seams). A statistical evaluation of the cured (sintered) compressive strengths proved that pellets prepared from UG2 ore are likely to have the same, or better, compressive strength than pellets prepared from metallurgical-grade chromite ore. The cured abrasion strength of the UG2 pellets was also superior to that of the metallurgical-grade pellets. Scanning electron microscopy (SEM) backscatter, secondary electron, and elemental X-ray mapping were used to determine the reasons for the general superior strength of the UG2 pellets. The case study UG2 ore also required 13 kWh/t less energy for milling to attain the required particle size distribution prior to pelletization, which can lead to substantial cost savings. Results presented in this paper can be utilized by ferrochromium (FeCr) producers to better quantify the advantages and disadvantages associated with the use of UG2 ore for FeCr production.

Keywords South Africa, UG2, metallurgical-grade chromite, oxidative sintered pellets, compressive and abrasion strengths, milling energy requirements.

Introduction Chromite, a mineral of the spinel group, is the only commercially viable source of virgin chromium units. Geologically, commercial chromite deposits in the world are found in three forms, i.e. alluvial-, podiform-, and stratiform-type deposits (Murthy et al., 2011; Cramer et al., 2004). Alluvial deposits were formed by weathering of chromite-bearing rock with the subsequent release of chromite and gravity concentration by flowing water.

However, these deposits are relatively small and of comparatively minor commercial interest. Podiform-type chromite deposits usually occur as irregularly shaped pods or lenses, and their distribution within a mineralized zone is usually relatively erratic and unpredictable, making the exploration of these deposits difficult and costly (Cramer et al., 2004). Stratiform-type chromite deposits occur as parallel seams in large, layered igneous rock complexes with more regular layering and lateral continuity (Cramer et al., 2004). The largest example of a stratiformtype chromite deposit is the Bushveld Complex The Journal of The Southern African Institute of Mining and Metallurgy

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* Chemical Resource Beneficiation, North-West University, Potchefstroom. † Minerals Processing Division, Mintek. ‡ Glencore Alloys, South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Nov. 2014 and revised paper received Feb. 2015.

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Synopsis

(BC) in South Africa, which holds an estimated three-quarters of the world’s viable chromite ore resources (Cramer et al., 2004).

Several chromite seams exist in the BC. The seams of economic interest are the Lower Group 6 (LG6), the Middle Group 1 and 2 (MG1 and 2), and the Upper Group 2 (UG2) seams (Cramer et al., 2004). The LG and MG seams are specifically exploited for their chromium content, while the UG2 seam is mined primarily as a source of platinum group minerals (PGMs) (Xiao and Laplante 2004).

Extraction of the PGMs from the UG2 ore usually involves the liberation of the sulphide by milling and subsequent recovery of the PGM concentrate through flotation (Xiao and Laplante, 2004). Chromite in the PGM concentrate is undesirable; therefore, PGM recovery circuits are specifically designed to ensure maximum rejection of the chromite to the tailings stream. However, the chromite in the PGM tailings can potentially serve as feedstock for ferrochrome (FeCr) production, after beneficiation to increase the chromium content.

South Africa has dominated global production of FeCr in the recent past (Kleynhans et al., 2012 and references therein) due to its vast chromite resources and past history of relatively inexpensive electricity. FeCr is a relatively crude alloy, consisting predominantly of iron (Fe) and chromium (Cr), which is utilized mainly for the production of stainless steel. In South Africa, FeCr is generally produced from metallurgicalgrade chromite ore originating from the LG or MG seams. However, due to the increase in the availability of UG2 ore, as well as technical innovations in the FeCr and stainless steel industries, an increasing volume of UG2 feed

Comparison of physical properties of oxidative sintered pellets material is being utilized by South African FeCr producers (Cramer et al., 2004). However, there are certain advantages and disadvantages that need to be considered. The price that a FeCr producer is prepared to pay for UG2 is determined by balancing these advantages and possible disadvantages.

Advantages related to the use of UG2 by South African FeCr producers include: ➤ The utilization of UG2 for FeCr production minimizes tailing volumes for PGM producers, which also has associated financial benefits, e.g. smaller tailing storage facilities ➤ The use of UG2 contributes to the optimum utilization of natural resources, since additional mining is avoided. This also reduces the carbon footprint of FeCr producers ➤ The overall safety risk associated with FeCr production is much lower if the mining of new chromite units can be reduced by the utilization of upgraded UG2 tailings from PGM operations. Although this is not a direct financial or technical benefit, environmental-, healthand safety-related aspects are as important as production volumes and profit margins in the modernday industrial setting ➤ UG2 ore is usually cheaper than conventional metallurgical chromite ore from LG6 or the MG1 and 2 seams.

The disadvantages associated with the substitution of metallurgical-grade chromite ore with UG2 during FeCr production include: ➤ The Cr/Fe ratio of UG2 ore is lower than that of metallurgical-grade ore from LG6 or the MG1 and 2 seams (Cramer et al., 2004). As well as Fe and Cr, chromite also contains magnesium (Mg) and aluminium (Al) in varying proportions, depending on the deposit. The chemical composition of chromite can be represented by the formula (Mg, Fe2+) (Cr, Al, Fe3+)2O4 (Murthy et al., 2011). In this structure, Fe2+ can be replaced by Mg and similarly Cr can be replaced

by Fe3+ and Al. This variation in composition can result in substantial differences in the Cr/Fe ratio of chromite deposits. Specifically within the BC, the compositional variation within a particular seam is relatively small across the whole Complex, but there is a progressive shift in composition between the sequential reef horizons. Since Fe oxides are more easily reduced (at lower temperatures) than Cr oxide (Niemelä et al., 2004), Fe recovery from chromite ore is higher than that of Cr. Therefore, the lower Cr/Fe ratio of UG2 ores results in a lower Cr content of the FeCr produced. Since FeCr producers are paid only for the Cr content of FeCr, UG2 utilization has a disadvantage in this regard ➤ The above-mentioned dilution of Cr content in the FeCr produced from UG2 also results in higher transport costs per Cr unit. This is significant, considering that most of South Africa’s FeCr is exported to stainless steel producers in Europe, Asia, and North America.

This additional transport cost also increases the overall carbon footprint of FeCr production in South Africa ➤ Cramer et al. (2004) reported that the higher Ti content of UG2 ore could also be problematic, since some stainless steel producers have specifications for the Ti content of FeCr. FeCr is produced in South Africa with various process combinations, which were recently reviewed by Beukes et al.

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(2010). At least eight of the 14 South African FeCr smelters utilize oxidative sintered pelletized feed. As described by Beukes et al. (op. cit.), this process entails the wet milling of chromite, together with a small percentage of carbonaceous material, followed by de-watering. Refind clay, which serves as a binder is then added and mixed into the moist milled material. The mixture is then pelletized in a pelletizing drum.

The over- and undersized green pellets are removed and recycled, while the appropriately-sized green pellets are layered on a sintering belt, which is protected by a layer of previously sintered pellets. The green pellets are then ignited in a furnace, and air is passed through the pellet bed to sinter the pellets. The amount of carbon present in the green pellets is limited to supply just enough exothermic energy to sinter the pellets properly (Beukes et al., 2010).

The oxidative sintered pelletized feed technology is commonly applied in the South African FeCr industry, and beneficiated UG2 ore is increasingly being used as a feedstock into this process. However, a direct comparison of the physical properties of oxidative sintered pellets produced from UG2 with the properties of pellets produced with conventional metallurgical-grade chromite ore (from the LG6 or the MG1 and 2 seams) is currently lacking in the peerreviewed scientific literature. Singh and Rao (2008) reported the impacts of properties of different chromite ores of Indian origin on pelletization and sintering. Although knowledge can be gained from similar studies, it cannot be used to exactly predict the performance of South African UG2 ore in the oxidative sintered process. This knowledge gap was recently highlighted by a large South African FeCr producer that requested the authors to investigate this matter. This producer indicated that some negative aspects associated with the physical properties of oxidative sintered pellets produced from UG2 ore were being reported by operational personnel. This had resulted in some operational resistance to the use of large proportions of ore from the UG2 horizon in the oxidative sintered process. In this paper, the physical properties of oxidative sintered pellets produced from a typical South African BC metallurgical-grade ore are compared to the properties of oxidative sintered pellets produced from a typical South African UG2 ore.

Experimental Materials A UG2 ore sample and a metallurgical-grade ore sample (from the LG6 seam) were received from one of the largest FeCr smelters situated in the western limb of the BC. At the time when this study was initiated, these specific ores were used for the production of oxidative sintered pellets at this smelter. Refined bentonite (G&W Base (Pty) Ltd., 2013), which was used as a binder during the pelletization process associated with the oxidative sinter pellet production, was obtained from the same FeCr producer. Anthracite from Nkomati Anthracite (Pty) Ltd, Mpumalanga Province, South Africa (Sentula, 2013) was used as a carbonaceous source in the pellets.

In addition, 10 milled chromite ore samples were obtained from a large FeCr producer applying the oxidative sintered process. These samples were used only to determine the particle size requirement during milling of chromite ore in preparation for the production of oxidative sintered pellets.

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Comparison of physical properties of oxidative sintered pellets Surface, crystalline, and chemical analyses Scanning electron microscopy (SEM) using energy-dispersive X-ray spectroscopy (EDS), utilizing an FEI Quanta 200 instrument with an integrated Oxford Instruments INCA 200 EDS microanalysis system, was utilized to perform surface analysis of the case study materials, i.e. visual inspection and surface chemical characterization. Chemical characterization of chromite ore samples was performed with a Spectro Ciros Vision inductively coupled plasma optical emission spectrometer (ICP-OES). Semi-quantitative and quantitative X-ray diffraction (XRD) analyses were conducted with a Philips X-ray diffractometer (PW 3040/60 X’Pert Pro) to determine the mineral composition of the UG2 and metallurgical-grade chromite ores. The samples were scanned using X-rays generated by a copper (Cu) Kα X-ray tube. The measurements were carried out between variable divergence and fixed receiving slits. The crystalline phases were identified using X’Pert Highscore Plus software, while the relative phase amounts were estimated using the Rietveld method (Autoquan programme). X-ray fluorescence (XRF) was also used to determine the concentration of elements present in the case study materials, using the same instrument as for XRD analysis, with a rhodium (Rh) X-ray tube and a Super Q database.

diffraction particle sizing using a Malvern Mastersizer 2000.

A diluted suspension of milled material was treated ultrasonically for 60 seconds prior to the measurement in order to disperse the individual particles and to avoid the use of a chemical dispersant.

Pelletization The milled material mixtures (obtained from the Siebtechnik mill) were pressed into cylindrical pellets in a Specac PT No. 3000 13 mm die with an LRX Plus strength-testing machine (Ametek Lloyd Instruments) equipped with a 5 kN load cell.

Each pellet consisted of 3 g dry milled material to which two drops of water were added prior to pressing the pellet, i.e. a moisture content of approximately 4.1 wt%. The compression rate was controlled at 10 mm/min until a load of 1.5 kN was reached, and this load was then held for 30 seconds.

Although time-consuming (each pellet was made individually), this technique was preferred over conventional disc or drum pelletization, which can result in the formation of pellets with different densities and sizes. Pressing the pellets according to the procedure described here ensured consistent density and size, allowing a monovariance investigation of other process parameters.

Milling

Oxidative sintering

A Siebtechnik mill was used to reduce the materials (ore, anthracite, and bentonite) to the required particle size prior to pelletization. All parts of this mill that made contact with the charge made of tungsten carbide, which prevented possible Fe contamination of the milled material. Prior to milling, the materials were dried at 40ºC for one day and then cooled in airtight containers to avoid possible moisture absorption. The three components were mixed in a ratio of 97 wt% ore, 2 wt% anthracite and 1 wt% bentonite. Batches of 170 g of this dried material mixture were then milled. After a batch had been milled, the milled material mixture was collected in a sample bag and stored for future usage. The mill was cleaned before milling of a new batch commenced.

In order to quantify the energy requirements during the milling of the two different case study ores, stirred mill tests were conducted. These were carried out utilizing a mill with a 5 L stationary vertical chamber and a stirrer consisting of a shaft fitted with pins to agitate the grinding media. The stirrer rotational speed was held constant at 400 r/min (3.7 m/s). The grinding chamber was 18 cm in diameter and 23 cm high. The mill was equipped with a variable speed drive and a torque sensor to facilitate the accurate determination of energy absorbed by the charge. All the key operating conditions were monitored with a computer that also calculated the energy consumption. The grinding chamber was equipped with a water jacket for cooling. 1 kg of ore was milled with 1.5 kg of water and 20 kg of 6 mm steel balls. For each case study ore, five different milling energy inputs were tested experimentally, i.e. 10, 20, 40, 80 and 160 kWh/t. In practice (industrial application), a relatively narrow energy input range would be expected. However, this relatively large range was tested in order to identify possible trends.

A camber furnace (Lenton Elite, UK, model BRF 15/5) with a programmable temperature controller was used to conduct all oxidative sintering experiments. The temperature profile used in this experimental study was compiled in an attempt to simulate conditions occurring in the industrial oxidative sintering process. Figure 1 depicts the temperature profile that was utilized. The furnace was heated up at the maximum rate to 1 400°C, and this temperature was maintained for 20 minutes. The furnace was then switched off and the pellets allowed to cool inside the furnace. After 100 minutes, the furnace door was opened to accelerate the cooling of the pellets.

Compressive strength testing The compressive strength of the oxidative sintered pellets was tested with an Ametek Lloyd Instruments LRXplus strength tester. The speed of the compression plates was maintained at 10 mm/min during crushing to apply an increasing force on the pellets. The maximum force applied to incur breakage was recorded for each pellet.

Particle size analysis

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Figure 1 – Temperature profile used to cure oxidative sintered pellets VOLUME 115

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Particle size distribution of the as-received case study ores was determined by wet screening. The particle size distributions of the milled chromite ore samples generated with the two different milling methods were determined by laser

Comparison of physical properties of oxidative sintered pellets Abrasion resistance testing The abrasion resistance test apparatus utilized was based on a downscaled version of the European standard EN 15051 rotating drum, similar that used by Kleynhans et al. (2012) and Neizel et al. (2013). The drum was designed according to specifications provided by Schneider and Jensen (2008).

Abrasion resistance tests were conducted at drum rotational speeds of 33, 55, and 77 r/min. A batch of ten oxidative sintered pellets, generated under specific experimental conditions, was abraded for different time periods, i.e. 1, 2, 4, 8, 16, 32, and 64 minutes. After each time interval, the material was screened at 9.5 mm. The over- and undersized material was then weighed and all the material returned to the drum for further abrasion until the final abrasion time interval was reached.

Statistical handling of data The compressive strength tests were repeated 50 times for each of the experimental conditions investigated. In order to determine whether the differences in mean values between the compressive strength of the two case study ores were meaningful, the t-test was applied (Skoog et al., 2004).

from the process residue of the PGM industry and had already been milled to liberate the PGMs. The relevance of the difference in size distribution of the as-received case study ores will be discussed later.

In Table II, the chemical, surface chemical, and crystalline compositions of the two case study ores are presented.

Chemical analysis (ICP-OES) indicated that the Cr2O3 contents of the metallurgical grade and UG2 case study ores were 44.19 and 41.82%, respectively, which is typical for these ores (Cramer et al., 2004). Slightly lower Cr2O3 contents were obtained for both case study ores with XRF analysis, which can be expected since results obtained from the destructive chemical analysis technique (ICP-OES) will be more accurate. As expected, the XRD analyses indicated that chromite was the dominant Cr-containing mineral phase. All chemical analytical techniques (ICP-OES, SEM-EDS, and

Table I

Size distribution of as-received case study ores wt.%

Results and discussion Material characterization The particle size distributions of the as-received case study UG2 and metallurgical-grade chromite ore samples are presented in Table I. It is evident from these results that the UG2 ore was substantially finer than the metallurgical-grade chromite ore. This was expected, since the UG2 ore originated

Particle size (Îźm)

Metallurgical-grade chromite ore

Beneficiated UG2 process residue

<90 90-106 106-150 150-300 300-600 >600 TOTAL

1.66 0.39 2.54 20.17 69.3 5.94 100

10.54 2.58 23.44 42.23 20.78 0.43 100

Table II

Characterization of the two case study ores in terms of the chemical (ICP-OES and XRF), surface chemical (SEM-EDS), and crystalline (XRD) contents XRF Na2O MgO Al2O3 SiO2 P2O5 SO3 Cl K2O CaO TiO2 Cr2O3 MnO Fe2O3 Co3O4 NiO ZnO ZrO2 ICP-OES Cr2O3 FeO SiO2 Al2O3 MgO CaO P Cr/Fe

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Metgrade 8.46 11.0 4.48 0.01 0.02 0.01 0.02 0.24 0.67 40.7 26.4 0.13 Metgrade 44.19 24.68 2.96 14.71 10.31 0.16 0.001 1.58

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UG2 0.12 9.1 11.7 5.46 0.01 0.03 0.01 0.01 0.35 0.78 37.8 0.28 27.3 0.09 0.17 0.0849 0.01 UG2 41.82 26.2 3.38 16.21 10.47 0.27 0.001 1.4

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XRD semi-quantitative Magnetite Chromite Enstatite Magnesioferrite Tetrataenite Trevorite SEM-EDS Cr Fe Al Mg Si Ca Ti V O

Metgrade 9 38 32 16 4 Metgrade 28.1 16.0 6.5 6.3 2.4 0.4 40.3

UG2 32 58 4 2 4 UG2 28.1 18.0 6.7 5.7 2.0 0.3 0.4 0.3 38.5

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Comparison of physical properties of oxidative sintered pellets

Particle size requirement For the pelletized chromite pre-reduction process, Kleynhans et al. (2012) stated that the milled particle size requirement is 90% of the particles (d90) to be smaller than 75 μm.

However, definitive information regarding the particle size requirement for oxidative sintered pellets could not be obtained in the peer-reviewed public domain. Laser diffraction particle size analyses of 10 milled chromite ore samples obtained from a large FeCr producer applying the oxidative sintered process yielded an average d90 of 108 μm.

This value was therefore adopted as the required particles size target during subsequent milling experiments.

Figure 2 presents the d10, d50, and d90 values obtained by laser diffraction particle size analysis of the two case study ores milled for different time periods with the Siebtechnik mill. The solid lines indicate the trends of each data series.

Although the d50 of both the case study ores was the same after approximately 95 seconds of milling, the d90 of the metallurgical-grade chromite ore at this milling time was still higher than the target value of 108 μm. After approximately 120 seconds of milling, both case study ores had a d90 close to the aforementioned target value. However, at this milling time, the d50 values differed substantially. This data indicated that it would be impossible with this specific mill to find a milling time that would yield a d90 value of 108 μm for both ores, with both ores also having the same particle size distri-

Figure 2 – Particle size of the two case study ores, milled with the Siebtechnik mill, used to prepare oxidative sintered pellets The Journal of The Southern African Institute of Mining and Metallurgy

bution. This might have been possible if the composition of the milling media could have been changed, but this was impossible with the mill utilized. It was therefore decided to mill the case study ore mixtures for both 95 and 120 seconds in preparation for pelletization.

Compressive strength comparison In Figure 3, the compressive strength of oxidative sintered pellets prepared from the two case study ores milled with the Siebtechnik mill for 95 seconds (i.e. with the same d50 values) and for 120 seconds (i.e. with similar d90 values), is presented. In this box-and-whisker plot, the short horizontal line inside the box indicates the median of each data-set, the dot the mean, the upper and lower ends of the boxes the 25th and 75th percentiles, and the whiskers 2.7σ (99.3% coverage if it is assumed that the data has a normal distribution).

Although there are significant overlaps between the datasets, it seems that the oxidative sintered pellets prepared from UG2 ore milled for 95 seconds were slightly stronger (i.e. higher median and mean values) in relation to the oxidative sintered pellets prepared from the metallurgicalgrade chromite ore milled for 95 seconds. The calculation of the t-test values and comparison of these to standard tvalues at a 95% probability (Skoog et al., 2004) proved that the differences in the mean pellet compressive strengths between these two data-sets were meaningful. However, for pellets prepared from ore milled for 120 seconds, there was no meaningful difference in the compressive strengths of the pellets. It can therefore be stated with a high level of confidence that oxidative sintered pellets prepared from UG2 ore are likely to have a similar or better compressive strength compared to pellets prepared from metallurgical-grade chromite ore.

Abrasion strength comparison The abrasion strengths at different rotational drum speeds for oxidative sintered pellets prepared from UG2 and metallurgical-grade chromite ore are presented in Figure 4.

All of these pellets were prepared from ore milled for 120 seconds in the Siebtechnik mill. It is evident from this data (Figure 4) that the abrasion strength of the UG2 oxidative sintered pellets was better than that of the oxidative sintered pellets prepared from metallurgical-grade chromite ore.

Figure 3 – Statistical comparison of compressive strength of oxidative sintered pellets prepared from the two case study ores milled for 95 seconds and 120 seconds with the Siebtechnik mill VOLUME 115

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XRF) indicated that the Fe content of the UG2 ore was higher than that of the metallurgical-grade chromite ore, which was reflected in the Cr/Fe ratios of 1.4 and 1.58 respectively.

These Cr/Fe ratios are quite typical of South African metallurgical-grade and UG2 ores that are used in FeCr production (Cramer et al., 2004).

The bentonite used as a binder and the anthracite used as a carbonaceous source in the oxidative sintered pellets were exactly the same sample materials utilized by Kleynhans et al. (2012). Kleynhans et al. conducted detailed characterization of these two materials (chemical, surface chemical, and crystalline contents) and this data is therefore not presented in this paper. In summary, the bentonite consisted mostly of smectite mineral groups, which were made up of montmorillonite. On an air-dried basis, the anthracite sample contained 75.08% fixed carbon, 17.79% ash, and 6.87% volatiles.

Comparison of physical properties of oxidative sintered pellets Microstructural analysis In order to assess why the compressive and abrasion strengths of oxidative sintered pellets prepared from UG2 ore were in general better than those of oxidative sintered pellets prepared from the metallurgical-grade chromite ore, microstructural analysis was performed. In Figure 5, SEM backscatter micrographs of polished sections of oxidative sintered pellets prepared from metallurgical-grade and UG2 chromite ore are presented at two magnifications. All these pellets were prepared from ore milled for 120 seconds with the Siebtechnik mill. From the micrographs, it is evident that the oxidative sintered pellets prepared from both case study ores had a crystalline structure. However, the pellets prepared from the metallurgical-grade chromite ore had a much finer and better defined crystalline structure than that of the pellets prepared from UG2 ore.

According to the Hall-Petch law (Equation [1]), structures with smaller grain size withstand a higher force before breaking (DoITPoMS, 2013).

[1]

Figure 4 – Abrasion resistance strength indicated in weight percentage remaining above 9.5 mm versus abrasion time

where σy is the tensile yield stress, d is the grain diameter, σi is the intrinsic yield stress, and k is a constant for a particular material.

If the Hall-Petch law is applied, the pellets prepared from metallurgical-grade chromite would be expected to be stronger. This was clearly not the case (Figures 3 and 4).

However, the Hall-Petch law applies only to grain structures of materials with the same composition. As indicated in Table II, the chemical and crystalline characteristics of the two case study ores differ. It is therefore unlikely that this law can be applied.

SEM secondary electron micrographs of the areas shown in Figure 5 are presented in Figure 6. In contrast to SEM backscatter micrographs (Figure 5), secondary electron micrographs (Figure 6) give a topographic perspective. Since these figures illustrate polished sections of both pellet types, limited depth perspective information can be derived from Figure 6. However, it seems as if the crystalline structure in the oxidative sintered pellets prepared from the metallurgicalgrade chromite is better defined (Figures 6(a) and 6(b)) than the crystalline structure in the pellets prepared from the UG2, which blends into the general matrix (Figures 6(c) and 6(d)).

This indicates that the crystals in the oxidative sintered pellets prepared from the metallurgical-grade chromite constituted a well-defined separate phase from the matrix, while in the pellets prepared from the UG2 ore, the crystals had assimilated better into the general matrix. The better assimilation of the crystalline growth into the matrix of the pellets prepared from the UG2 ore could at least partially explain why this type of pellet was stronger than that prepared form metallurgical-grade chromite (Figure 3 and 4).

X-ray mapping of elemental distribution in the two pellet types is presented in Figure 7. It is evident that the Mg distribution correlated best with the crystalline structures observed (Figures 5 and 6), with low Mg content in the crystals

Figure 5 – Backscatter SEM micrographs, at two different magnifications, of oxidative sintered pellets prepared from metallurgical-grade chromite ore (a and b) and from UG2 ore (c and d)

Figure 6 – SEM secondary electron micrographs, at two different magnifications, of oxidative sintered pellets prepared from metallurgical-grade chromite ore (a and b) and from UG2 ore (c and d)

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Comparison of physical properties of oxidative sintered pellets

Figure 7 – SEM elemental X-ray mapping micrographs of oxidative sintered pellets prepared from metallurgical-grade chromite ore and from UG2 ore

utions, and Fe and Mg distribution can therefore not be used to explain the differences in pellet strengths observed (Figures 3 and 4). Further research should be undertaken to investigate the possibly link between elemental distribution and the strength of oxidative sintered pellets prepared from metallurgical-grade chromite and UG2 ores.

Milling energy requirements

compared to the general matrix. Although not that welldefined, Fe was enriched in the crystals observed. Kapure et al. (2010) also observed Fe, in the form of the sequioxide phase (Fe2O3), precipitated on the surface of chromite grains in a typical Widmanstatten pattern, after pre-oxidation of chromite ore. However, in the current study both types of oxidative sintered pellet exhibited similar elemental distribThe Journal of The Southern African Institute of Mining and Metallurgy

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Figure 8 – Energy consumption as a function of particle size during stirred mill experiments for the two case study ores. The solid horizontal line indicates the target d90 value of 108 μm

In Figure 8, the energy input measured during the stirred mill tests is presented as a function of particle size, i.e. d90 and d50, for both case study ores. The targeted particle size, i.e. d90 of 108 μm, is indicated as a solid horizontal line. From these results, it is evident that the case study UG2 ore required 13 kWh/t less energy than the metallurgical-grade chromite to reach the target particle size. This is plausible, since the particle size distribution of the as-received UG2 ore was substantially finer than that of the as-received metallurgical-grade chromite (Table I). Although Cramer et al. (2004) mentioned that less energy is required to mill UG2 ore than metallurgical-grade ore, this advantage has not yet been quantified in the public peer-reviewed domain. With electricity costs rising rapidly in South Africa (Kleynhans et al., 2012) and electricity being the single largest cost component in FeCr production (Daavittila et al., 2004), this possible electricity saving during milling prior to pelletization could be significant for large FeCr producers. For example, if

Comparison of physical properties of oxidative sintered pellets a generic FeCr producer with a production capacity of 300 000 t FeCr per annum is considered and it is assumed that 2.1 t of oxidative sintered pellets are consumed to produce 1 t FeCr, it can be calculated that approximately 4.15 million South African rands can be saved if the case study UG2 ore was used exclusively instead of the case study metallurgical-grade chromite ore. This calculation is based on an electricity unit cost of R0.523 per kWh (Eskom, 2011).

DAAVITTILA, J., HONKANIEMI, M., and JOKINEN, P. 2004. The transformation of ferrochromium smelting technologies during the last decades. Journal of the South African Institute of Mining and Metallurgy, October. pp. 541–549.

DoITPoMS. 2013. University of Cambridge. http://www.doitpoms.ac.uk/tlplib/mechanical-testing/grainsize.php [Accessed 7 July 2013].

Conclusions The results presented for the case study ores prove that the compressive strength of pellets prepared from the UG2 ore was the same, or better, than that of pellets prepared from the metallurgical-grade chromite. The UG2 pellets also had superior abrasion strength. It can therefore be stated that the use of UG2 ore, instead of metallurgical-grade chromite ore, is unlikely to result in the formation of additional fines (e.g. less than 6 mm) during material handling and feeding into submerged arc furnaces (SAFs). Only limited fine materials can be accommodated in SAFs, since fines can lead to unstable operation, as well as increased safety risks and equipment damage. SEM backscatter, as well as secondary electron and elemental X-ray mapping, indicated that the cured oxidative sintered pellets prepared from both the case study ores had developed a relatively well-defined crystalline structure. The crystalline structure of the pellets prepared from the UG2 ore was, however, better assimilated into the matrix, which could explain its general superior strength. The results further prove that less energy was required to mill the case study UG2 to the desired particle size, which can lead to substantial cost savings.

The results of this investigation can be used by South African FeCr producers to better quantify the advantages or disadvantages associated with using UG2 ore. Considering the size of the South African FeCr industry, decisions impacting the local FeCr industry also have an influence on the global FeCr market. Additionally, substantial volumes of UG2 ore are currently being exported for FeCr production outside South Africa. Therefore, the results can also assist these producers to contextualize the use of South African UG2 ore.

ESKOM. 2011. Eskom retail tariff adjustment for 2011/2012. http://www.eskom.co.za/content/priceincrease2011.pdf [Accessed 13 October 2011].

G&W BASE (PTY) LTD. 2013. Bentonite Ocean MD. http://www.gwbase.co.za/pdfs/Pel-Bond.pdf [Accessed 16 July 2013].

KAPURE, G., TATHAVADKAR, V., RAO, C.B., RAO, S.M., and RAJU, K.S. 2010. Coal based direct reduction of peroxidised chromite ore at high temperature.

Proceedings of the 12th International Ferro-Alloys Congress (INFACON XII). Helsinki, Finland, 6–9 June 2010. pp. 293–301.

KLEYNHANS, E.L.J., BEUKES, J.P, VAN ZYL, P.G, KESTENS, P.H.I., and LANGA, J.M. 2012. Unique challenges of clay binders in a pelletised chromite prereduction process. Minerals Engineering, vol. 34. pp. 55–62.

MURTHY, Y.R., TRIPATHY, S.K., and KUMAR, C.R. 2011. Chrome ore beneficiation challenges and opportunities – a review. Minerals Engineering, vol. 24, no.

5. pp. 375–380.

NIEMELÄ, P., KROGERUS, H., and OIKARINEN, P. 2004. Formation, characterization and utilization of CO-gas formed in ferrochrome smelting. INFACON X, Proceedings of the Tenth International Ferroalloys Congress: Transformation through Technology, Cape Town, South Africa, 1-4 February 2004. South African Institute of Mining and Metallurgy, Johannesburg. pp. 68–77.

NEIZEL, B.W., BEUKES, J.P, VAN ZYL, P.G, and DAWSON, N.F. 2013. Why is CaCO3 not used as an additive in the pelletised chromite pre-reduction process?

Minerals Engineering, vol. 45. pp. 115–120.

Acknowledgements The authors wish to thank Glencore Alloys for financial assistance and Mintek for the use of their stirred mill.

SCHNEIDER, T. and JENSEN, K.A. 2008. Combined single-drop and rotating drum dustiness test of fine to nanosize powders. Annals of Occupational Hygiene, vol. 52, no. 1. pp. 23–34.

References SENTULA. 2013. Nkomati Anthracite (Pty) Ltd. http://www.sentula.co.za BEUKES, J.P., DAWSON, N.F., and VAN ZYL, P.G. 2010. Theoretical and practical aspects of Cr(VI) in the South African ferrochrome industry. Journal of the Southern African Institute of Mining and Metallurgy, vol. 110. pp. 743–750.

[Accessed 18 July2013].

SINGH, V. and RAO, S.M. 2008. Study of the effect of chromite ore properties on pelletisation process. International Journal of Mineral Processing, vol. 88.

BEUKES, J.P., VAN ZYL, P.G., and RAS, M. 2012. Treatment of Cr(VI) containing wastes in the South African ferrochrome industry - a review of currently applied methods. Journal of the Southern African Institute of Mining and Metallurgy, vol. 112. pp. 347–352.

CRAMER L.A., BASSON, J., and NELSON, L.R. 2004. The impact of platinum production from UG2 ore on ferrochrome production in South Africa.

Journal of the South African Institute of Mining and Metallurgy, vol. 104, no. 9. pp 517–527.

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SKOOG, D.A, WEST, D.M, HOLLER, F.J., and CROUCH, S.R. 2004. Fundamentals of Analytical Chemistry. 8th edn. (International Student Edition). Thomson Learning, Brooks/Cole. 155 pp.

XIAO, Z. and LAPLANTE, A.R. 2004. Characterizing and recovering the platinum

group minerals – a review. Minerals Engineering, vol. 17. pp. 961–979. ◆ The Journal of The Southern African Institute of Mining and Metallurgy

http://dx.doi.org/10.17159/2411-9717/2015/v115n8a7

The influence of selected biomass additions on the co-pyrolysis with an inertinite-rich medium rank C grade South African coal

by C.A. Strydom*, T.Z. Sehume*, J.R. Bunt*†, and J.C. van Dyk*‡

Co-pyrolysis of four biomass samples (hardwood chip, softwood chip, pinewood chip, and sugarcane bagasse) with an inertinite-rich medium rank C South African coal was investigated. Proximate and ultimate analyses of the chars prepared using a heating rate of 10°C/min up to 1100°C in a nitrogen atmosphere were used to compare the properties of the biomass and the coal chars. Similar gross calorific values (28.5–29.1 MJ/kg) for the woody biomass-coal blended chars were observed, which were slightly higher than that of the coal char sample (26.5 MJ/kg). CO2 surface areas of the chars of the woody biomass samples (328.1–329.4 m2/g) and of the blends (238.5–271.5 m2/g) were higher than that of the coal char sample (94.2 m2/g). Thermogravimetric (TG), differential thermal analyses (DTG), and calculated weighted averaged TG curves indicated that the influence of the biomass on the pyrolysis rate of the coal is small and vice versa. The CO2-producing reactions of the coal were slightly enhanced during co-pyrolysis.

Keywords inertinite-rich coal, biomass, co-pyrolysis, TG-MS, CO2 evolution

Introduction Coal utilization has led to rising concerns about adverse impacts on the environment (global warming) caused by toxic gases (H2S, SOx, NOx, and CO2) and remaining waste.

Biomass is considered to be CO2-neutral with regard to the greenhouse gas balance, and is regarded as a renewable source that assists in reducing CO2 emissions when compared with coal (McGowan, 1991; McKendry, 2002).

Biomass currently makes up approximately 14% of the world’s energy sources (McGowan, 1991; McKendry, 2002). Coal combustion and gasification are prominent processes in South Africa (in producing synthesis gas and energy) due to the abundance of coal in the country.

Some of South Africa’s coal resources are, however, high in ash content and also are inertinite-rich, which renders them more difficult to use industrially than vitrinite-rich low-ash coal (Jeffrey, 2005; Strydom et al., 2011). There is a need to investigate the properties and behaviour of the inertinite-rich high-ash South African coals and especially their interaction with local biomass sources.

The gasification process occurs through various stages, of which pyrolysis of the char residue forms an integral part (Bunt and The Journal of The Southern African Institute of Mining and Metallurgy

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Materials and methods Coal and biomass samples An inertinite-rich medium rank C bituminous coal from the Witbank coalfield in South Africa, previously analysed in detail by Hattingh et al. (2013), was obtained. The sample was milled and sieved to a size fraction of -75 μm and stored under nitrogen. Hattingh et al. (2013) described the analysis techniques

* Chemical Resource Beneficiation, North-West University, Potchefstroom, South Africa. † Sasol Technology (Pty) Ltd, South Africa. ‡ African Carbon Energy, South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Oct. 2014 and revised paper received Feb. 2015.

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Synopsis

Waanders, 2009). Currently, particular interest is shown in the co-utilization of coal and biomass to produce synthesis gas via the gasification process (Usón et al., 2010; McGowan, 1991; McKendry, 2002).

Investigating the co-pyrolysis of the South African coal and local biomass material is thus the first step in evaluating the use of local biomass sources for co-gasification purposes.

No reported research on the influence of biomass during co-pyrolysis with high-ash inertinite-rich coal is available. Previous studies focused on fast pyrolysis of biomass and coal-biomass blends to produce bio-oil, as well as co-gasification using vitrinite-rich coals and biomass (Collot et al., 1999; Moghtaderi et al., 2004; Kumabe et al., 2007; De Jong et al., 1999). A better understanding of lowmolecular-mass gaseous products formed by the pyrolysis process during co-utilization is required. In this work, the co-pyrolysis behaviour of the original materials (coal, hardwood chips, softwood chips, pinewood chips, and sugarcane bagasse) and their blends with an inertinite-rich South African coal is reported. Properties of the resulting chars were also investigated and will be discussed.

The influence of selected biomass additions on the co-pyrolysis and characteristics of the coal sample in detail. The coal sample contains 24.4% vitrinite, 72.2% inertinite, and 3.3% liptinite (mineral-matter-free basis). The vitrinite reflectance is reported as 0.81 Rr% (Hattingh et al., 2013). The sample contains 83.8% carbon, 4.3% hydrogen, 11.5% oxygen, 2.1% nitrogen, and 1.0% sulphur (DAF basis), and the proximate analysis indicated 15.2% ash, 25.2% volatile matter, and 59.6% fixed carbon (DB).

The biomass used was chosen based on the seasonal availability of resources. The Sappi Forest Division Company (South Africa) supplied the following biomass materials: (1) soft woodchip (SWC), (2) hard woodchip (HWC), and (3) pine woodchip (PWC). TSB Sugar Company (South Africa) supplied sugarcane bagasse (SB). The biomass samples were pulverized to -75 μm. Prior to further investigation, all samples were dried in an oven at 80°C for 24 hours to remove surface moisture. All samples were mixed thoroughly using a small ball mill before further use, and were stored in a desiccator flushed with N2. Samples were characterized using the standard methods for chemical and mineralogical analysis, as summarized in Table I. XRF analyses were performed according to the ASTM D4326 standard method (ASTM D’4326, 2012). A Micromeritics ASAP 2010 analyser was used to determine the surface areas of the char samples using CO2 adsorption.

Table I

Standard chemical and mineralogical analysis methods used to analyse samples Standard method

Pyrolysis reactivity A known weight percentage of biomass (0%, 20%, 40%, 60%, 80%, and 100%) was added to the inertinite-rich coal and mechanically mixed using a Wig-l-Bug. The pyrolysis behaviour of the coal, biomass, and their blended samples were investigated using a SDTQ 600 thermogravimetric analyser coupled to a Cirrus MKS quadruple mass spectrometer. Sample masses of approximately 25 mg of coal, biomass, or of the blended samples were thermally treated in a nitrogen environment from ambient temperature to 1100°C, using a heating rate of 10°C/min. Al2O3 ceramic sample pans were used. The evolution of some of the low-molecular-mass gaseous species (H2, CH4, H2O, and CO2) was recorded simultaneously with the mass loss data. The mass loss and gaseous product curves were recorded as a function of temperature and time during the thermal treatment of each sample. Data acquisition was repeated at least three times to ensure repeatability. Averaged curves were used for further analysis.

Larger amounts of the coal, biomass, and their blended chars were also prepared in a tube furnace. Ceramic pans (150 mm × 45 mm × 12 mm) were used to load a known sample mass into the centre of the tube furnace. The system was then flushed with N2 for 10 minutes prior to heat treatment. The heating programme used to prepare the chars was identical to the heat treatment of the samples in the thermogravimetric analyser. After reaching 1100ºC, the samples were cooled to room temperature in the tube furnace under the nitrogen atmosphere (duration of cooling time was approximately 6 hours).

Results and discussion Char properties

Proximate analysis

Ultimate analysis

Inherent moisture (%) Ash yield (%) Volatile matter (%) Fixed carbon (%) Carbon, hydrogen, nitrogen, sulphur (%)

Calorific value

SANS ISO 5925 SABS ISO 1171 SABS ISO 562 By difference ISO 12902

The ultimate and proximate analysis data and CO2 surface areas of the chars of the coal, biomass, and blended samples, are summarized in Table II. The high ash content of the sugarcane bagasse char and 80% SB + 20% coal char could be due to soil particles having been incorporated during harvesting. The high ash content of

ISO 1928

Table II

Ultimate and proximate analyses of the char samples prepared from the coal, biomass and 80% biomass + 20% coal blended samples (heating rate 10°C/min to 1100°C in N2) Proximate analysis (DB*)

Inherent moisture (wt.%) Ash (wt.%) Volatiles (wt.%) Fixed carbon (wt.%) Gross calorific value (MJ/kg) Ultimate analysis (DAF**) Carbon (wt.%) Hydrogen (wt.%) Nitrogen (wt.%) Total sulphur (wt.%) Oxygen (wt.%) CO2 BET surface area (m2/g) * Dry basis; ** Dry ash-free basis

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Coal char

HWC char

SWC char

PWC char

SB char

80% HWC + 20% coal char

80% SWC + 20% coal char

80% PWC + 20% coal char

80% SB + 20% coal char

2.4 18.3 1.8 77.5 26.5

2.9 0.9 2.4 93.8 28.6

2.8 0.9 1.8 94.5 30.8

3.1 1.0 2.4 93.5 28.6

2.9 45.8 1.9 49.4 17.4

3.4 10.7 1.2 84.7 28.9

3.1 9.5 1.3 86.1 28.5

2.8 9.0 1.9 86.3 29.1

2.6 50.8 0.7 45.9 17.0

97.2 0.1 1.1 0.4 1.2 94.2

94.5 0.1 0.4 0.0 5.0 328.1

93.2 0.2 0.6 0.0 6.0 332.4

95.1 0.1 0.5 0.0 4.3 329.4

89.1 0.2 0.3 0.3 10.1 100.5

97.2 0.2 0.7 0.4 1.5 271.5

95.7 0.0 0.7 0.2 3.4 256.5

93.2 0.2 0.8 0.2 5.6 238.5

92.8 0.2 0.7 0.2 6.1 107.3

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The influence of selected biomass additions on the co-pyrolysis the SB sample also resulted in a lower calorific value, as expected. The gross calorific values of the biomass chars were higher than that of the coal char, except for the SB biomass sample, which contains a very high percentage of ash. The gross calorific values of the blended samples (28.5–29.1 MJ/kg) are close to that of the coal and biomass char samples (26.5–30.8 MJ/kg ), except for the SB biomass sample, which as indicated has a high ash content. When harvesting biomass for use during co-processing with coal, care should be taken not to include soil in the biomass samples. Co-pyrolysis of coal and biomass seems to result in a similar product (char) in terms of calorific value.

Ultimate analysis results suggest that the SB biomass sample reacts differently from the woody biomass samples, as less carbon and more oxygen are retained in the biomass char sample. The surface area of the SB biomass char sample was also much lower than the values for the woody biomass chars, but this could also be due to the large amount of ash that the SB char contains.

The surface areas of the biomass chars were greater than those of the coal and blended chars. The surface areas of the blended sample chars were observed to be still much higher than that of the coal char, indicating that the pores of biomass chars were not substantially blocked by carbonaceous deposits (Darmstadt et al., 2001).

XRF analysis data for the charred samples of the coal, biomass, and 80% biomass+20% coal blended samples are given in Table III. XRF analysis of the SB shows that the sample contains relatively large amounts of Si and Fe, measured as SiO2 and Fe2O3, indicating the presence of soil, as also evident from the measured ash percentage (Table II).

respectively, with SWC showing the lowest residue percentage (15 wt. %). The mass loss curves of the coal sample indicated approximately 75% unreacted mass (char).

Figure 2 presents the derivative thermogravimetric (DTG) curves of the biomass samples and of the coal. The curves show that the degradation range of the biomass samples is approximately 180–525°C. The PWC and SWC samples show

Figure 1 – Mass loss data with increasing temperature for parent materials (heating rate 10°C/min and N2 atmosphere)

Thermogravimetric analysis Figure 1 shows the curves of mass loss versus temperature for the coal and four biomass samples heated at 10°C/min under a nitrogen atmosphere in the TGA. The mass loss curves indicate that SB has the largest amount of residue char (35 wt.%) of all the biomass samples after heating to 1100°C. The larger mass percentage residue obtained is due to the high ash content of SB. The unreacted mass percentages of PWC and HWC were 17 wt.% and 16 wt.%

Figure 2 – DTG curves of parent materials (heating rate 10°C/min and N2 atmosphere)

Table III

XRF analysis data of the chars of coal, biomass, and 80% biomass + 20% coal blended samples (wt.% of total inorganic species) HWC char

SiO2 54.2 58.0 Al2O3 28.2 29.1 Fe2O3 1.8 2.3 TiO2 1.4 1.3 P2O5 1.5 0.3 CaO 4.3 1.9 MgO 1.0 1.1 Na2O 0.1 0.2 K2O 0.6 1.0 SO3 6.2 3.0 Unidentified 0.8 1.8 * All blends are 80% biomass + 20% coal.

SWC char

PWC char

SB char

HWC + coal char*

SWC + coal char*

PWC + coal char*

SB + coal char*

54.5 30.6 2.5 1.2 0.3 1.1 1.1 0.9 1.2 5.8 0.9

53.7 30.2 1.3 1.3 0.4 2.5 1.0 0.2 0.9 6.9 1.6

56.2 25.4 10.1 1.2 1.3 0.9 1.1 0.9 1.6 0.2 1.1

49.0 29.6 5.2 1.5 1.5 5.3 0.9 0.1 1.1 6.8 0.9

52.2 29.5 7.3 1.9 0.5 2.2 1.5 1.0 1.1 2.0 0.9

50.2 32.5 1.5 1.3 1.6 4.4 1.4 0.2 1.0 4.6 1.3

51.9 29.5 4.2 2.1 1.4 3.6 1.9 0.5 0.9 3.7 0.3

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Coal char

The influence of selected biomass additions on the co-pyrolysis a single peak with maximum mass loss rate at approximately 360°C for both biomasses. The SB and HWC samples exhibit two major peaks, which overlap and have maximum rates of mass loss at 308 and 353°C (SB) and 289 and 357°C (HWC).

The well-described main composition of biomass is hemicellulose, cellulose, and lignin (Yang et al., 2007; Demirbas, 2000). The first peak is associated with the decomposition of hemicellulose, while the second peak (at the higher temperature) is associated with the decomposition of cellulose (Oudia et al., 2007; Demirbas, 2000). The four biomass samples exhibit shoulders on the DTG curves at approximately 440°C. These shoulder peaks are ascribed to the degradation of lignin, which occurs at a higher temperature than the decomposition of hemicellulose and

cellulose (Yang et al., 2007; Oudia et al., 2007; Demirbas, 2000).

The coal sample showed two distinctive DTG peaks at approximately 451°C and 695°C. According to Serio et al.

(1987) the release of primary volatiles (such as light hydrocarbons and condensable tars) from coal predominates at approximately 500°C. Radovic et al. (1983) stated that coal pyrolysis above 650°C was due to gases being released during condensation reactions, thermal decomposition of carbonates, and breakdown of aromatic rings.

Figure 3 presents the thermogravimetric (TG) and differential thermal analyses (DTG) curves of the coal and the blended samples. As expected, the higher the amount of biomass in the blend, the lower the char yield. The DTG

Figure 3 – TG and DTG curves of coal and blended coal-biomass samples. (a) HWC, (c) SWC, (e) PWC, and (g) SB. DTG curves of (b) HWC, (d) SWC, (f) PWC and (h) SB (heating rate 10°C/min and N2 atmosphere)

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The influence of selected biomass additions on the co-pyrolysis Table IV

DTG peak maximum temperatures during pyrolysis as determined from DTG curves (heating rate 10°C/min and N2 atmosphere) Coal peak 1 maximum t (°C)

Coal peak 2 maximum t (°C)

362 365 364 363 363 357 362 360 359 358 360 366 365 364 363 353 360 358 357 357

451 448 440 437

695 679 678 675 670 680 679 671 667 700 700 677 672 669 667 666 644

curves of the blended samples are similar to those of the individual materials. The devolatilization process of the blended samples started at approximately 160°C and was completed at around 900°C. Some of the DTG peak maximum temperatures are listed in Table IV to facilitate investigation of possible shifts in the maximum rate of reaction steps.

The DTG peak maximum values for the degradation step at approximately 360°C do not differ substantially (353–365°C) and no trend was observed. The DTG peak maximum values of the coal devolatilization step at approximately 450°C also only differ between 437 and 451°C although, where visible, the DTG peak maxima for the blends were lower than that of the coal sample. The blended samples’ DTG peak maximum values, just lower than 700°C, show a decreasing trend with an increase in biomass percentage in the blends. The values are also lower than that for coal, except for the PWC sample, but then the value did not differ substantially (+5°C). These trends will be compared to the evolution of gases as observed from the mass spectroscopic data.

The effect of the four selected biomass samples on the pyrolysis of coal and the possible effect of the coal on the pyrolysis of the biomass samples were investigated. This was done by calculating the theoretical mass loss curves (weighted averages) from the TG curves of the raw materials.

The theoretical mass loss curves were calculated using the following equation: Ycalculated = x1Y1 + x2Y2 where x1 and x2 represent the fractions of the coal and biomass in the blends, and Y1 and Y2 are the mass loss percentages of coal and biomass, respectively, at the various temperatures. The calculated weighted average TG curves were compared to the experimentally derived results.

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449 445 441 449 444 437 445 445 445

Figure 4 presents a comparison between the experimentally obtained conversions and calculated mass loss curves for the 80% biomass + 20% coal blended samples. The curves for all the other blends follow the same trends.

The deviation at each temperature was calculated using the following equation: Deviation% = (Experimental remaining mass%−Calulated remaining mass%)/Experimental remaining mass% The average deviation percentage of each curve was calculated. The average deviation and observed maximum deviation percentages (with the temperature range in which the maximum deviation was observed) are listed in Table V. The calculated weighted average TG curves for all the blends were close to the experimentally determined results, with a maximum average deviation of less than 3% and a maximum deviation percentage of less than 12% observed for the 80% SWC + 20% coal blend at approximately 360°C. The maximum deviation percentages for the higher biomass–tocoal ratio blends occurred at approximately 360°C under the experimental conditions and were pronounced only for HWC and SWC. The decomposition of cellulose occurs at approximately 360°C (Yang et al., 2007; Oudia et al., 2007; Demirbas, 2000).

These small average differences between the calculated TG curves and the experimental TG curves indicate that the influence of the biomass samples on the coal reactivity is small, and vice versa. These results are similar to those obtained by Moghtaderi et al. (2004), Sonobe et al. (2008), and Sadhukhan et al. (2008) using blended samples of coal and Radiata pine wood, corncob, and waste wood biomass samples.

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Coal SWC 20% SWC + 80% coal 40% SWC + 60% coal 60% SWC + 40% coal 80% SWC + 20% coal HWC 20% HWC + 80% coal 40% HWC + 60% coal 60% HWC + 40% coal 80% HWC + 20% coal PWC 20% PWC + 80% coal 40% PWC + 60% coal 60% PWC + 40% coal 80% PWC + 20% coal SB 20% SB + 80% coal 40% SB + 60% coal 60% SB + 40% coal 80% SB + 20% coal

Cellulose peak maximum t (°C)

The influence of selected biomass additions on the co-pyrolysis

Figure 4 – Experimental TG curves for blended samples compared with calculated weighted average TG curves for (a) 80% SWC + 20% coal, (b) 80% PWC + 20% coal, (c) 80% HWC + 20% coal, and (d) 80% SB + 20% coal (heating rate 10°C/min and N2 atmosphere)

Table V

Average deviation and observed maximum deviation (with temperature ranges in which the maximum deviations were observed) for differences between calculated TG curves and experimentally obtained TG curves Blends

Blending ratio* (%)

Average deviation (%)

Maximum deviation (%)

t, °C (maximum deviation %)

0.3 1.0 0.0 2.8 0.7 0.0 1.0 0.8 0.3 1.0 1.0 0.3 0.6 1.0 1.0 1.0

1.7 2.0 3.0 10.8 1.8 2.0 8.0 11.3 1.6 2.0 1.0 1.2 1.4 3.0 1.0 4.0

356–360 373–513 377–409 368–369 1102 351–369 362 363 367–369 1061–1092 381–489, 514–1027 392–407 1089–1091 349-368 328–1100 353–365

HWC-coal

20–80% 40–60% 60–40% 80–20% SWC-coal 20–80% 40–60% 60–40% 80–20% PWC-coal 20–80% 40–60% 60–40% 80–20% SB-coal 20–80% 40–60% 60–40% 80–20% * Blend ratios are given as %biomass–%coal

Mass spectroscopy To obtain a better understanding of the influence of copyrolysis of biomass and coal on the process, the mass spectroscopic data for gases H2, CH4, H2O, and CO2, as they were evolved at the various temperatures during the pyrolysis of the samples, were investigated.

Figure 5 shows the evolution profiles of H2 versus temperature at a heating rate of 10°C/min under an inert

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environment for biomass samples, coal, and blends. The evolution of H2 was followed by observing the H2+ mass spectroscopic data (m/z = 2). The evolution of H2 increased with an increase in temperature, as was expected, and a maximum rate of H2 evolution is reached for all samples at approximately 740°C, except for the SB biomass sample, which showed a maximum rate of H2 evolution at approximately 680°C. The H2 evolution profiles of the blends stayed similar in shape with an increase of the blending ratio.

The Journal of The Southern African Institute of Mining and Metallurgy

The influence of selected biomass additions on the co-pyrolysis

Figure 5 – Evolution curves of H2 during pyrolysis. (a) SWC and coal, (b) PWC and coal, (c) HWC and coal, and (d) SB and coal (heating rate 10°C/min and N2 atmosphere)

The Journal of The Southern African Institute of Mining and Metallurgy

the DTG curves (Figure 3). The CH4 temperature evolution profiles of the blends are similar in shape and have similar peak maximum temperature values to the starting materials.

The biomass thus does not seem to influence the pyrolysis degradation reactions evolving CH4 in coal and vice versa.

Figure 7 shows the evolution of H2O with increasing temperature during the pyrolysis of coal, biomass, and blended samples. Moisture is released below 150°C from all the samples. The four biomass samples started evolving H2O around 200°C in at least two overlapping steps (as also observed in the DTG curves in Figure 3). The evolution of H2O in this temperature range is related to the thermal decomposition of hemicellulose and cellulose in the biomass samples (Bassilakis et al., 2001; Worasuwannarak et al., 2007; Huang et al., 2011; Demirbas, 2000).

H2O evolution from the coal started at approximately 320°C, reaching a maximum rate at 520°C. This is due to condensation reactions occurring and can be explained by the relatively high oxygen content in the coal sample (Di Nola et al., 2010; Gavalas et al. 1981). Production of H2O from all the blends was between 250°C and 500°C. The H2O evolution profiles of the blends between 200°C and 400°C stayed similar in shape with an increase in blending ratio. The H2O temperature evolution profiles of the blends are similar in shape and have similar peak maximum temperature values to the starting materials. The biomass thus does not seem to influence the pyrolysis degradation reactions evolving H2O in coal and vice versa.

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The chemical mechanisms of evolution of H2 during devolatilization of coal and biomass are described in the literature (Yang et al., 2007; Van Heek and Hodek, 1994; Demirbas, 2000; Gavalas et al., 1981). The similar shapes and peak maximum temperature values indicate that the reactions by which H2 is evolved during devolatilization of the blends are not significantly changed by adding biomass to coal. The biomass thus does not seem to influence the pyrolysis degradation reactions evolving H2 in coal, and vice versa.

Figure 6 illustrates the evolution of CH4 with an increase in temperature during the devolatilization of coal, biomass, and blended samples. The fragment CH3+, with m/z value of 15, was chosen to represent methane (CH4) in order to prevent interference from oxygen (m/z of 16). Methane is known to easily ionize by giving off hydrogen ions in the mass spectrometer (Huang et al., 2011). From Figure 6 it can be observed that the formation of CH4 started at temperatures just above 210°C for biomass and blended samples, and the evolution of CH4 from coal started at temperatures above 400°C. A similar evolution of CH4 was observed by other authors for the thermal degradation of coal during pyrolysis, and the chemical reactions where methane is evolved during pyrolysis of coal have been well described (Yang et al., 2007; Van Heek and Hodek, 1994; Di Nola et al., 2010; Gavalas et al., 1981).

The evolution of methane from the biomass and blended samples during pyrolysis showed a similar pattern to that of

The influence of selected biomass additions on the co-pyrolysis

Figure 6 – Evolution curves of CH4 (measured as CH3+ at m/z=15) during pyrolysis of (a) SWC and coal, (b) PWC and coal, (c) HWC and coal, and (d) SB and coal (heating rate 10°C/min and N2 atmosphere)

Figure 7 – Evolution curves of H2O during pyrolysis of (a) SWC and coal, (b) PWC and coal, (c) HWC and coal, and (d) SB and coal (heating rate 10°C/min and N2 atmosphere)

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Figure 8 – CO2 evolution curves during pyrolysis of (a) SWC and coal, (b) PWC and coal, (c) HWC and coal, and (d) SB and coal (heating rate 10°C/min and N2 atmosphere)

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in peak maximum temperatures that are shifted lower to approximately 660°C for the blended samples. These results are confirmed by the DTG peak maximum values for the blends, which show a decreasing trend with increasing biomass percentage in the blends (Figure 3 and Table IV).

The decrease of approximately 35°C of the temperature at the maximum rate of CO2 evolution from the coal is significant, and indicates that the presence of biomass influences the chemical reactions through which CO2 is evolved from the coal. The CO2-producing reactions in coal seem to be catalysed by the presence of the biomass or the biomass degradation products, which could be either in the gaseous or the solid phase. The enhancement of the CO2 evolution reactions of coal is, however, only slight.

During co-pyrolysis with the biomass material, the inertinite-rich bituminous coal behaves similarly to the vitrinite-rich bituminous coals studied previously, except for the slight enhancement of the CO2 evolution reactions in the case of the inertinite-rich bituminous coal (Collo, et al., 1999; Moghtaderi et al., 2004; Kumabe et al., 2007; De Jong et al., 1999).

Conclusion Co-pyrolysis of the woody biomass samples and the inertinite-rich coal at a heating rate 10°C/min up to 1100°C resulted in chars with similar gross calorific values and carbon, nitrogen, and volatile mass percentages. TG, DTG, and MS results at the maximum rate of CO2 evolution showed VOLUME 115

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Figure 8 shows the evolution of CO2 with increasing temperature during the thermal degradation of all the samples in the N2 atmosphere. At lower temperatures, the evolution of CO2 during thermal degradation of a bituminous coal is due to the decomposition of aliphatic, aromatic carboxyl, and carboxylate groups (Gavals et al., 1981; Di Nola et al., 2010). At higher temperatures, CO2 evolution during thermal treatment is ascribed to the decomposition of the more stable ether structures, oxygen-bearing heterocyclic compounds, and carbonates. In addition, the existence of intra-molecular carboxylic acid anhydrides in the bituminous coal may also contribute to the formation of CO2 during pyrolysis (Gavalas et al., 1981; Van Heek and Hodek, 1994).

The CO2 evolution spectra of coal exhibited a single peak that started from approximately 400°C and reached maximum rates of CO2 evolution at 695°C. CO2 evolution from the biomass samples is derived mainly from the cracking and reforming of functional groups of carboxyl (C=O) and COOH in the biomass material (Yang et al., 2007). The spectra exhibited CO2 peaks that started at approximately 220°C for the biomass and blended samples and reached maximum peak heights at approximately 370°C. Below 500°C, the peak shapes and peak maximum temperature values for the blended samples are similar to those of the related biomass samples, and the presence of coal seems thus not to influence the thermal degradation of the biomass samples during pyrolysis.

The evolution of CO2 from the coal in the blends resulted

The influence of selected biomass additions on the co-pyrolysis that the co-pyrolysis slightly enhances the CO2-producing reactions in the coal. The reactions producing H2, CH4, and H2O were not influenced by co-pyrolysis. Co-pyrolysis of softwood chip, hardwood chip, pinewood chip, and sugarcane bagasse with the inertinite-rich rank C bituminous coal results in chars with slightly enhanced properties for gasification.

KUMABE, K., HANAOKA, T., FUJIMOTO, S., MINOWA, T., and SAKANISHI, K. 2007. Co-

Acknowledgements

MCKENDRY, P. 2002. Energy production from biomass (part 1): overview of

The authors would like to thank Sasol Technology Research and Development and North-West University for partial funding of this research. The work presented in this paper is based on research supported by the South African Research Chairs Initiative of the Department of Science and Technology and National Research Foundation of South Africa (Coal Research Chair Grant No. 86880).

gasification of woody biomass and coal with air and steam. Fuel, vol. 86, no. 5. pp. 684–689.

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http://dx.doi.org/10.17159/2411-9717/2015/v115n8a8

Mine Occupational Safety and Health Leading Practice Adoption System (MOSH) examined – the promise and pitfalls of this employer-led initiative to improve health and safety in South African Mines

by M. Hermanus*, N. Coulson*, and N. Pillay† summarizes the results of the evaluation and includes further analysis and discussion of the nature of the programme and the conditions that act to advance it or it hold back.

Synopsis This paper assesses the effectiveness of the Mine Occupational Health and Safety Leading Practice Adoption System (MOSH) and its potential to improve mine health and safety in South African mines. Developed by the Chamber of Mines, which represents the majority of the country's large scale mining employers, MOSH was devised to accelerate progress towards achieving health and safety milestones, which were set by tripartite agreement in 2003. The paper documents and builds on the findings of a study conducted by the Centre for Sustainability in Mining and Industry (CSMI) in 2011 that evaluated MOSH strategy, structures, and process of implementation.

The study found that MOSH operated across the mining sector, was directed and dominated by experts and, despite best efforts to include other stakeholders, was led by employers. Statutory worker health and safety representatives and structures were not integrated into the complex change process developed by MOSH. The Mine Health and Safety Inspectorate (MHSI) and organized labour were ambivalent about direct involvement in MOSH and preferred regulatory measures to enforce the participation of mines. MOSH interventions were not targeted at mines with a poor health and safety record, and MOSH lacked a baseline from which to track impacts on sector-wide health and safety performance. The leading practices most widely adopted by mines were designed to improve, rather than fundamentally alter, existing practice. Although the depth of engagement with MOSH among stakeholders and on mine sites varied, mining companies, labour representatives, and the Mine Health and Safety Inspectorate (MHSI) saw the programme as significant.

Background

Introduction In 2011, the Chamber of Mines commissioned the Centre for Sustainability in Mining and Industry (CSMI) to evaluate the progress and effectiveness of the rollout of the Mine Occupational Health and Safety Leading Practice Adoption System (MOSH). CSMI, based in the School of Mining Engineering at the University of the Witwatersrand, is an independent research centre conducting health and safety research relevant to all major stakeholders in the mining sector. The purpose of the evaluation was to identify the factors that enabled or hampered the adoption of new technologies or practices. The timing of the evaluation, two years before the deadline for the 2013 safety and health milestones, allowed time for adjustments to be made. This paper The Journal of The Southern African Institute of Mining and Metallurgy

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* Centre for Sustainability in Mining and Industry University of Witwatersrand. † Independent Consultant. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received May 2014 and revised paper received Dec. 2014.

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Keywords health and safety, regulatory framework, best practice, milestones.

Health and safety in the South African mining sector is governed by the Mine Health and Safety Act (MHSA). The MHSA reflects an outcomes approach to regulation, based on risk assessment and risk management. It provides for extensive dialogue and consultation between the Mine Health and Safety Inspectorate (MHSI), which represents government, employers and organized labour on matters of policy, regulation, research, and the state of health and safety in the sector. The MHSA also authorizes the MHSI to enforce the provisions of the Act through issuing directives to take corrective action, halting activities that pose imminent and high levels of risk to health and safety, and administering fines. In addition, the MHSI is charged with the responsibilities of promoting health and safety through investigations and inquiring into the causes of accidents, initiating research, making health and safety data public, and issuing advice.

Tripartite engagement is formalized in the Mine Health and Safety Council (MHSC). This is a national structure established by the MHSA that considers the state of health and safety in the sector, proposes policy and legislation, commissions research, and provides advice to the Minister of Mineral Resources. Through the MHSC, organized labour, government, and employers come together to undertake their mandated joint

Mine Occupational Safety and Health Leading Practice Adoption System (MOSH) examined responsibilities, which include a biennial review of health and safety. The review takes the form of a convention called the Mine Health and Safety Summit. At the Summit of 2003, the tripartite partners agreed to establish targets for health and safety, and intermediate milestones for the period 2003 to 2013 (DMR, 2013) (Table I). The 2010 Summit launched a framework that included guiding principles, commitments and action points, to shift health and safety culture, referred to in the sector as the Culture Transformation Framework.

As seen in Table I, the goals for 2013 involved reducing the number of fatalities to levels attained by the mining sectors of Australia, the USA, and the Ontario province of Canada; and from 2013 onwards, eliminating new cases of silicosis and noise-induced hearing loss. The attainment of these milestones represented a considerable challenge.

Before MOSH began piloting leading practices to address health and safety challenges in 2008, the sector relied solely on the individual efforts of mining companies to achieve the milestones. MOSH represented a shift to collective and informed effort, active promotion of promising interventions, and engagement of the entire sector, across all commodities and all regions of the country.

The MOSH Leading Practice Adoption System MOSH centred on the uptake of leading practices to address health and safety priorities and included both improved technology and procedures. The practices addressed four challenges in keeping with 2013 health and safety milestones, namely falls of ground; transport and machinery, relevant to the reduction of mine fatalities; and dust and noise, relevant to the elimination of silicosis and noiseinduced hearing loss respectively. Table II summarizes seven leading practices promoted through MOSH at the time of the CSMI process evaluation.

Leading practices promoted by MOSH were already established in industry, not yet widely used across the sector, and judged to offer significant health and safety hazard

mitigation potential. Identifying leading practice was the first step. The site at which a leading practice was first found was called the ‘source’ mine. The next major step involved piloting the leading practice in a second mine site, referred to as the ‘demonstration’ mine. To support piloting at the demonstration mine, research into the knowledge, attitudes, and values of workers, supervisors, and managers responsible for implementation was conducted, to inform subsequently developed behavioural communication and leadership plans. This research was referred to as understanding the ‘mental models’ of workers, supervisors, and management. The adoption process at the demonstration mine site was carefully documented and recommendations made about the process of adoption based on this experience.

Once successfully demonstrated, the leading practice was then actively promoted for adoption across the industry as a whole.

The MOSH leading practice adoption system was systematically documented in a handbook which described the structures required for driving the programme, provided tools and advice for overseeing implementation, and documented processes for conducting workshops and other key activities.

Other materials were also developed to support the adoption process, which included documentation on the MOSH portal, DVDs, and brochures.

The organizational structures of MOSH The organizational structures for MOSH in 2011 (and which were still in operation at the time of writing of this paper) included Adoption Teams, a MOSH Task Force, a Learning Hub, and communities of practice. These structures provide strategic and operational focus, enable learning within the system, and guide and encourage implementation at mine sites. At the time of the evaluation the structures were constituted as follows: ➤ MOSH Adoption Teams existed for each area of leading practice aligned with the four priority health

Table I

2003 health and safety targets and milestones (DMR, 2011; Pienaar and du Plessis, 2009; van der Woude et al., 2004) Target: zero rate of fatalities and injuries Milestone In the gold sector: by 2013, achieve safety performance levels equivalent to current international benchmarks for underground metalliferous mines at least.

In the platinum, coal, and other sectors: by 2013 achieve constant and continuous improvement equivalent to current international benchmarks at least.

Target: elimination of silicosis Milestones

By December 2008, 95% of all exposure measurement results will be below the occupational exposure limit for respirable crystalline silica of 0.1 mg/m3 (these results are individual readings and not average results).

After December 2013, using current diagnostic techniques, no new cases of silicosis will occur amongst previously unexposed individuals (previously unexposed individuals are individuals unexposed prior to 2008, i.e. equivalent to a new person entering the industry at 2008).

Target: elimination of noise-induced hearing loss (NIHL) The present noise exposure limit specified in regulation is 85 dB(A) Milestones After December 2008, the hearing conservation programme implemented by industry must ensure that there is no deterioration in hearing greater than 10% amongst occupationally exposed individuals.

By December 2013, the total noise emitted by all equipment installed in any workplace must not exceed a sound pressure level of 110 dB(A) at any location in that workplace (including individual pieces of equipment).

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Mine Occupational Safety and Health Leading Practice Adoption System (MOSH) examined Table II

Summary of the seven leading practices promoted by MOSH in 2011 Leading practice

Summary

Quieter rock-drill

THE HILTI ROCK DRILL TE MD20 is a quieter rock-drill to reduce noise exposure in underground mines.

Even though the Hilti rock-drill is much quieter than pneumatic rock-drills, it must be used together with an effective hearing protection device (HPD) if noise exposure is to be reduced to a level below the occupational exposure limit of 85 dBA. The Fogger is an installation to capture airborne respiratory dust. This is an atomized water dustsuppression system. It reduces the risk of airborne respirable crystalline silica (RCS) dust and other dust through engineering out the hazard. It offers dust control at the points where dust is generated. Once operational, it reduces the exposure of a large number of employees when applied together with other silica dust controls.

This is a dust suppression technique to reduce airborne dust in underground mines. It is a technique that involves wetting the footwall and sidewall with water and surfactants by means of a spray car or a hand pump.

In trackbound mines, spray cars are pulled by locomotives and as the solution is sprayed dust that has settled on the foot and sidewalls is coagulated. This prevents the dust from becoming airborne. The treatment must be repeated at appropriate time intervals to be effective. The frequency of treatment depends on the operating environment, and is determined at the mine site.

A method for reducing fatalities and injuries caused by falls of ground (FoG) in underground mines.

Bolts and nets are required and are installed in the hangingwall in the stoping areas of underground mines. To be effective the installation must: • Be as close to the hangingwall as possible and attached to mechanical props at the face • As close to the face as possible • Wide enough to provide coverage for most of the activities in the panel • Be held in place with roofbolts installed in a regular pattern • Have multiple attachment points to reduce the span of unsupported net • Where temporarily installed, be cleared of rockfalls, removed before the blast, and stored away from the blast area.

A selection and education package aimed at reducing noise exposure.

The tool is in two parts. The first supports the selection of HPDs and the second provides educational material to encourage compliance. The selection tool utilizes Excel™ software. It enables occupational hygienists on mines to choose the correct HPD for workers in different occupations based on noise exposure. Noise and frequency measurements for all occupations in the mining industry are presented, together with information on all the HPDs available in South Africa and the purpose for which they are fit. The educational package aims to raise awareness of the hazard, and encourage the wearing of appropriate (but at times uncomfortable) HPDs. It includes two DVDs, training manuals for trainers, and comic-strip booklets. The booklets are available in four languages.

Preventing moving machinery from injuring people and collisions between vehicles. A proximity detection system employs devices that sense the presence of nearby objects or people without needing to make physical contact to do so. Many types of devices are available and these use different means of signalling, such as wi-fi (radio waves), magnetic fields, or sonar for different ’targets’. The systems promoted by the MOSH leading practice initiative include devices for use in: • Hard rock mines with railbound equipment. These have loco–to-loco devices that give both visual and audible warning to the operator to slow down, and if the operator does not respond, the device stops the locomotive • Hard rock mine with trackless equipment. These have machine–to-machine OR machine–to-people devices, which give audible warnings • Soft rock mines with trackless equipment. These have machine–to-people devices, which warn, slow down the machinery/vehicle, and stop it if necessary.

A process of entry examination and making safe of current underground working places in order to minimize and eliminate fall-of-ground accidents Entry examination and making safe procedures are legally required to be carried out regularly in underground mines, on the basis of a risk assessment, typically at the start of a new shift entering a working place. This practice is critical to safety in stope panels and development ends, but in and of itself carries high risk.

Footwall and sidewall treatment

Nets with bolts

Hearing protection device (HPD) selection tool and training and awareness

Proximity detection device

Entry examination and making safe.

and safety challenges. These were led by one or two technical specialists, often seconded to the Chamber of Mines by industry for a minimum of six months (at the time of the evaluation). Rollout of leading practice across the industry proceeded under the leadership of the Adoption Teams ➤ The MOSH Task Force provided governance and oversaw the adoption activity. High-level industry The Journal of The Southern African Institute of Mining and Metallurgy

representatives served on the MOSH Task Force. The responsibilities included ensuring that Adoption Teams were well aligned to industry needs, and that the needs of the teams were communicated to industry ➤ The MOSH Learning Hub, at the Chamber, was responsible for managing the MOSH Adoption Teams to maximize industry ownership, exposure, and buy-in ➤ Communities of practice were ad hoc structures that VOLUME 115

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Fogger dust suppression system

Mine Occupational Safety and Health Leading Practice Adoption System (MOSH) examined met to disseminate information on leading practices.

Management, technical, and health and safety staff from mine sites were invited to attend a specific community of practice (also known as COPA) for a specific adoption practice. COPAs were facilitated in mining regions and often involved participants from across different mining companies and operations learning together.

Methodology of the evaluation study Leading practices were largely piloted in demonstration mines in 2008 and rolled out across the industry from 2009 onwards. Although MOSH was designed to increase uptake though the active promotion and guided adoption of leading practices, the programme progressed slowly and unevenly, prompting the evaluation study conducted by CSMI. The evaluation was conducted between August and November 2011 by a multidisciplinary team of researchers from CSMI, including researchers with backgrounds in mining engineering and geology, health and safety regulation, public health, and sociology. The team included an experienced researcher fluent in indigenous languages.

The evaluation was a three-tiered study of the MOSH intervention. Data was collected and organized under three main themes: (i) oversight and governance, (ii) resourcing and capacity, and (iii) operational or mine-site level intervention. The study involved the collection of qualitative data through key informant interviews and focus group discussions. Data was collected from all levels of the MOSH intervention from senior industry leaders and technical experts through to teams of employees involved with the implementation of specific adoption methodologies.

Underground/on-site visits at operations were also conducted to explore the specific context for leading practice adoption.

The evaluation included an extensive review of MOSH documentation, including reports, minutes of meetings, and education and communication tools. In total fifteen people were interviewed about oversight and governance, and twenty individuals were interviewed about resourcing and capacity. The majority of these interviews involved officials and representatives from industry and the Chamber of Mines with direct responsibility for MOSH. Representatives from organized labour and the MHSI were also engaged for their perspectives on the governance of MOSH. Eight site visits to mines were conducted involving discussions with 119 individuals. Of this number, a total of 55 workers were

reached through focus group discussions.

Three of the seven leading practices promoted by MOSH were selected for in-depth investigation by the research team.

This was because they represented leading practices where there was greater uptake, documentation, and access to sites that could be followed from source to adoption mines. Table III summarizes in generic terms the selection of mines for participation in the evaluation. However, it also shows that at the time of the evaluation, only two of these practices actually followed the MOSH process to conclusion, from the source mine to the adoption mine.

The key evaluation questions and sub-questions considered by the research team are listed below.

How is the process of MOSH adoption happening at a site level? With a focus on: ➤ Understanding the MOSH process as it unfolds in practice from the source mine through to implementation at the adoption mine ➤ Assessing the key factors which facilitate and/or limit adoption ➤ Understanding how the process is currently monitored and evaluated ➤ Understanding the rollout of leading practices across industry using the adoption teams.

Assess the appropriateness and success of the existing MOSH structures, with a focus on: ➤ Assessing the performance to date of the Learning Hub, other key role-players, and other component structures ➤ Evaluating the governance and accountability structures of the Learning Hub ➤ Understanding the Learning Hub’s relationships with key stakeholders and other bodies in the implementation of the MOSH leading practice system ➤ Understanding the division of labour between all the key role-players with regard to the implementation of the MOSH adoption system. ➤ Reviewing the capacity of the Learning Hub, other key role-players, and other component structures within a changing mining environment for meeting the MOSH mandate in future.

A limitation of the study was that the sample of leading practices examined largely represented the success stories of the MOSH leading practice adoption system. The motivation

Table III

Mines selected for the evaluation of leading practice Leading practice

Source mine

Demonstration mine

Adoption mine

Dust: foggers Noise: Hilti rockdrill

Gold mine 1 Gold mine 3

Gold mine 2 Gold mine 4

Falls of ground Entry examination and making safe

Platinum mine 1

Gold mine 5

Coal mine 1 Not applicable (Hilti rockdrill was not rolled out beyond the demonstration mine) Coal mine 2

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Mine Occupational Safety and Health Leading Practice Adoption System (MOSH) examined and resources at participating mine sites may not be representative of the sector as whole. Mines that were not participating in the MOSH intervention were not included in the evaluation and hence their accounts are absent from the findings and analysis.

Findings of the evaluation study The findings of the evaluation point to MOSH success at a mine-site level with some serious gaps at other levels of the programme, including in some cases mine-site level gaps. The findings are presented under five headings, with quotes taken from the research interviews to illustrate the points made.

Success at mine-site level relied on interpretation and a customized fit Successes at mine-site level were found to rely on both the ability of Adoption Team Leaders, as well as mine management to interpret, distil and tune the MOSH Adoption System for site-specific conditions (see Table IV).

At the level of mine site implementation, when mines had successfully committed to the adoption process there was often genuine engagement with the MOSH process. At the mines visited it was clear that management had invested significant resources in meeting the implementation requirements. This involved: ➤ Addressing problems arising during implementation ➤ Seeking, listening to, and acting on feedback ➤ Being prepared to go back to the drawing board when initial efforts failed ➤ Collecting data to support whether a leading practice was working or not ➤ With the exception of one site included in the evaluation visits, the communities of practice were described as a useful structure through which to share experiences.

It was at the level of the mine-site that the expertise and wisdom of the Adoption Team Leaders were widely appreciated. One of the key findings at mine-site level was that workers, supervisors, and mine management accepted and embraced the MOSH leading practices.

Responses to particular leading practices varied across mine sites as indicated in the accounts in Table IV. Certain leading practices were much easier to introduce. Entry examination and making safe far outstripped the progress made with other leading practices with respect to the number of mines adopting a leading practice. Investments in technology such as the Fogger dust suppression system or proximity detection systems presented much more complex challenges, both in terms of capital outlay and alignment with existing systems of risk management.

A gap in the MOSH process at the time of the evaluation was that MOSH did not provide guidance on when and how to engage with suppliers and original equipment manufacturers (OEMs). This was found to be particularly problematic when only one or two suppliers are available for a particular technology (‘… there are not enough technicians to attend to its failures’) and when technologies are being transferred across commodities and the supplier or OEMs are working out of their normal context.

The MOSH process was overly technical and complex The overly technical and complex process of adoption implementation attracted a fair measure of criticism. The 48 steps of the MOSH adoption process captured in a complex handbook that includes 16 sections and 52 two worksheets and reference documents resulted in many individuals, even those intimately involved with the intervention, being unable to state clearly ‘what is MOSH?’ The various elements and steps clouded the central objective as the statements below attest.

Table IV

Successes and lessons of leading practices at mine sites

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Evaluation report: case studies Making safe - implementer perspectives ‘Up to now it has been a positive experience. Since the adoption of this leading practice we have less accidents, not just fall of ground accidents but a reduction in other types of accidents as well. We have developed a better understanding between the workers and supervision, and they realize that we are serious about preventing accidents and looking after the workers’ wellbeing’ (miner and shift boss).

‘Previously we would not all go into the workplace and do an examination. Only a miner would go into the workplace and the whole gang would be left behind at the waiting place. But right now we all participate and there is a difference. We can all pick up things that need to be fixed. We would fix things … we think it is a good system. … because we meet as a team and advise each other when we see something wrong.’ (Underground work team).

The Hilti TE MD20 electric drill - what did the implementers learn?

The evaluators noted that the introduction of the Hilti drill as a means to address excessive exposure to noise could not be supported if a compelling economic case could not also be made. This was the rationale at the source mine, where the Hilti rockdrill was introduced primarily for economic reasons.

As the continued operation of the mine depended on the Hilti, everyone on the mine had an interest in seeing the change in drilling technology succeed. The fact that management and workers freely shared information and concerns also helped. The direct benefits of the Hilti included reduced incidence of NIHL, a reduction of 40% in compensation payable, reduced shift losses, and cost savings of R67 434.00 per panel (2008 costs).

Attempts to transfer the Hilti rockdrill to mines where problems in production did not dictate the need for a change in drilling technology were not successful. Commitments to solving the problem of noise alone proved insufficient to disrupt well-established drilling practice, especially as the change also introduced new risks associated with the weight of the Hilti drill and the speed of start-up.

The Fogger - what did the implementers learn?

The Fogger had the potential to reduce exposure to dust. The source mine reported a decreasing respirable dust trend and occupational exposure levels (OELs) of silica dust consistently below 0.1 mg/m3. At the demonstration mine, by the end of the project 59% of employees were aware that the fogger unit was making a difference to their lives. The transfer of the technology to the demonstration coal mines presented new challenges due to different mining conditions to those associated with gold or platinum mines, at which Foggers were first employed.

Mine Occupational Safety and Health Leading Practice Adoption System (MOSH) examined ‘It’s too many steps and the challenge is to make it much simpler. Mining people are simple people – we don’t need laborious procedures. Decide what we are going to do and we do it and measure our progress. We are adding too much complexity to a simple matter.’

‘The management steps can be put in a nutshell. We did this and did not fixate on a piece of paper.’

The complex process also meant that common-sense steps in a change management process, such as ‘buy-in’ for a new initiative, did not happen optimally. Rather, ‘engagement’ tended to be codified in the process of understanding ‘mental models’ and the preparation of ‘communication and leadership plans.’

‘Is it possible to use other terms than mental models “getting understanding of how people think” or “perception survey”?’

Workers and other role-players felt marginalized by the MOSH process Codification also meant that labour representatives and other actors felt left out or excluded from important activities.

‘With hindsight we should have involved the engineering and health/medical sections of our mine in the MOSH project.’

‘I would say buy-in from underground teams is critical.

This was made clear to me in the failure in … leading practice. There is sometimes a fine line between consultation and buy-in. A pilot shouldn’t be in one area but rather in a shaft or across a level because this is a better platform to roll out from. I don’t have a recommendation to the MOSH process other than don’t rush it and don’t force feed … we followed the steps but we did not use the templates. Actually we used a mass meeting to kick-start the whole process.’

On paper, a strength of the MOSH process is its commitment to worker involvement; however, in practice this commitment and the activities designed to foster involvement (such as mental models referenced above) did not always translate into a whole-hearted experience of participation.

Workers were prepared for the introduction of changes through MOSH through a combination of demonstration, one or more days of training, and on-the-job coaching. Workers did not recognize the informal on–the-job coaching as training.

‘Demonstration for one day! You can’t call that training!’

‘We train ourselves … I talk to my team mate.’

Evaluation respondents used the term ‘resistance’ fairly loosely to describe all the concerns, fears, and objections to the MOSH process. The basis for this resistance could be well-founded. For example, leading practices can interfere with the pressure to meet production targets. Entry examination and making safe can and did result in longer periods of time spent securing the work area at the start of a shift, as the following quote from an underground supervisor indicated.

‘Although we were not very supportive at the introduction stage of the practice, as it affected our production rate, we can see the benefits now: it gives a clear message to the workforce that we care for their safety. The unions and associations are supportive.’

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There was an absence of a cogent strategy The MOSH Task Team, constituted by industry representatives, was considered as the primary structure involved with MOSH governance.

Respondents reported that MOSH did well in breaking down traditional hierarchical and inter-company barriers and helped to raise the profile of occupational health and safety issues in the workplace. However, elsewhere in the evaluation findings it was also reported that the biggest obstacle to widespread support for MOSH was the ‘not invented here’ mentality that results in disinclination at lower levels to implement the strategy.

‘The silo mentality is starting to disappear; we are in this as an industry together.’

‘At least the CEOs are starting to sing from the same hymn sheet – we face the same challenges and are starting to get solutions.’

Given the scope of the programme and the tasks involved, the evaluation team found, surprisingly, that MOSH lacked a cogent strategy. Dissimilar ideas about the intention(s) of MOSH surfaced in the responses of individuals at different levels of the programme to the question, ‘what is MOSH?’ At the operational site level, MOSH practitioners talked about ’finding leading practice’ or ‘sharing leading practice.’ In contrast, role-players at the level of governance and oversight placed more emphasis on the change management component of MOSH. ‘MOSH is a comprehensive change management system.’

‘MOSH changes the occupational health and safety culture in the industry.’

Although most respondents agreed that the MOSH initiative was in part a response to the 2013 milestones and that the impact of MOSH should be benchmarked against them, no hard success measures were applied to MOSH at the time of the evaluation. The ‘number of mines adopting’ was the most frequently quoted measure of MOSH progress, with little or no reference to the limitations of such a measure.

Other potential measures quoted by interviewees were reducing fatalities, creating a single industry standard, enhancing the skills of workers, and serving as a culture transformation framework.

The absence of a well-articulated strategy also meant that the Adoption Teams were faced with issues that they were unable to resolve. For example, it was acknowledged by the Dust Adoption Team that some of the leading practice technologies promoted through MOSH sat low down in the hierarchy of controls and posed a dilemma for practitioners who understand first principles. Over-simplifying the strategy or target setting for MOSH could result in leading practices being taken off-the-shelf rather than being an integral component of a well-reasoned health and safety risk management system at a specific mine site. If the adoption of leading practice was not based on risk management principles, MOSH could be accused of simply ‘selling’ leading practice. The inclusion of MOSH in the amended 2010 Mining Charter (refer to the Discussion section) was cited as an illustration of this problem.

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MOSH was industry- and expert-led rather than led by tripartite partners While MOSH made provision for the input of government and labour representatives at all stages of the MOSH process, such involvement was patchy and particularly superficial in planning workshops and decision-making structures of the programme. This underpined the observation that the Chamber of Mine largely drove the programme on its own.

Consequently, technical expertise and employer leadership characterized the MOSH intervention, and this steered the evaluation team to describe the intervention as ‘expert-led.’

‘As labour we were brought in very late. Labour needed to be part of the research process.’ ‘… has been searching for experts, a range of experts who can bring different expert dimensions to bear.’

‘If a person is not a recognized leader in the industry then you can’t succeed.’

The MHSI expressed ambivalence about MOSH for a number of reasons. First, it would prefer to engage with the programme though the MHSC, and secondly it was uncomfortable with the idea of promoting practices low down the hierarchy of controls. The MHSI also felt that the Chamber of Mines should act against its members with the worst health and safety performance. The fact that the Chamber of Mines is not representative of all mines was also a challenge to the MHSI. Organized labour regarded the impact of MOSH as significant at some mine-site levels but negligible in the mining sector as a whole. Trade unions regarded the involvement of workers in making informed decisions as an important outcome of MOSH. Labour felt they had an important role to play at all levels of the MOSH intervention, but recognized their limited capacity to do this. Like the The Journal of The Southern African Institute of Mining and Metallurgy

MHSI, they also felt unable to resolve and articulate their own role within MOSH and that the MHSC is the correct place for strategic discussions about MOSH.

Discussion What is the ‘promise’ MOSH represented? Evidence at a site level is that MOSH leading practices would be transferred between different sites, and where successfully transferred, could have beneficial effects. However, as shown in the evaluation, not all practices were easily replicated on different mines, and relevant leading practices or technologies were not necessarily taken up where they were most needed. This meant that the potential of MOSH was not fully realized. The MOSH evaluation study found that the programme did not target specific mines where improvements in health and safety performance would be of most significance for the entire sector. Thus the fundamental logic that runs through MOSH – change your health and safety culture, introduce effective technology and procedures, and improve the sectors’ health and safety performance – seemed to falter at the first post.

This begs the question as to whether there were gaps in the underlying or prevailing wisdom that were pertinent to the prospects of MOSH as a premier vehicle for improving health and safety. This is particularly important in the light of the new health and safety milestones that were set for the South African industry in 2014. In addition, what other considerations would make MOSH worthy of continued support and investment of time and resources?

The discussion here explores these questions by considering three areas that represent both the promise and the pitfalls of MOSH. These are the contribution of MOSH in the present climate of health and safety improvement; the exercise of leadership for MOSH within the health and safety regulatory framework; and readiness for change and worker participation in MOSH at the mine-site level.

The contribution of MOSH to sector-wide health and safety performance The data in Table V summarizes the fatality data at the start, mid-term, and the year preceding the end of the period for which the milestones apply. It was estimated that the safety performance of the South African mining industry, measured as fatality rate per million hours worked, must have improved by at least 20% per year to reach the average combined fatality rates of Australia, the USA, and Ontario (Canada) by 2013, (DMR, 2011; van der Woude et al., 2004,

Table V

Fatality rate per million hours worked – South African mining vs. average combined rates for mining in Australia, the USA, and Canada (Ontario) (Msiza, 2013) Year

Significance

Actual

Benchmark

2003 2008 2012

Initiative starts MOSH Leading Practice starts Penultimate milestone year

0.30 0.15 0.10

0.07 0.05 0.03

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‘Mining Charter both a driver and now a threat (compliance thereof). Not based on risk management but expectations of implementing a leading practice.’

The lack of cogent strategy also meant that decisions about targeting priority environments and/or mining commodities for leading practice were not made. The Transport and Machinery Adoption Team had an overwhelming task demonstrating the proximity device in three different setting: coal and hard rock trackless and hard rock railbound machinery.

‘We felt people will respond negatively to us if we just select hard rock and railbound transport, although this is where the number of fatalities are found – we haven’t targeted our intervention.’

It was striking how little power the existing structure felt it had to influence strategy. MOSH governance could be concluded to be ‘everybody else’s business.’ Problem companies with poor health and safety records could not be targeted by the existing MOSH structures; a point that is returned to in the discussion below.

‘There is no process for targeting problem companies.’

The most constructive suggestions regarding the way forward were that the tripartite MHSC should develop strategy and that the 2013 milestones could be used as a basis to set further targets for the next decade.

Mine Occupational Safety and Health Leading Practice Adoption System (MOSH) examined p. 6). While the South African mining sector had significantly closed the gap between itself and the benchmark countries in terms of fatality rates, the actual number of fatalities remained high, at 112 in 2012; and the year-on-year improvements were more modest than planned. The most significant reductions in the actual numbers of fatalities were achieved on coal mines (-30%) followed by gold mines (-18%).

Figure 1 shows how the rate of fatalities in South African mines improved over time relative to the benchmark over the period 2003 to 2012 (Msiza, 2013).

Fatality figures were the starting point for the risk assessments that illustrated the importance of specific leading practices. Using fatality figures from 2009 for falls of ground, it was estimated that appropriate interventions could save up to 129 lives per year. However, although the goal of MOSH is to impact on industry-wide health and safety performance, the improvements that have been achieved across the sector have not been attributed (other than in general terms) by the Chamber of Mines or the MHSI to any specific interventions promoted through MOSH, such as netting and bolting safety systems (Creamer, 2013; DMR, 2012).

Data of similar detail was not available for health, but progress was described as follows. Exposures to crystalline silica reported to the MHSI were all less than the occupational exposure limit of 0.01 mg/m3, but cases of silicosis continued to be diagnosed among miners. Whether the target of no new cases among previously unexposed individuals was achieved was unclear. In 2012, 1075 miners were diagnosed with noise-induced hearing loss of occupational origin, and the largest numbers were associated with gold and platinum mining (422 and 368 respectively). According to the MHSI, there was ‘no significant improvement on occupational disease’ (Msiza, 2013).

The difficulty of attributing safety improvements to MOSH was compounded by the fact that other major developments were unfolding at the same time, such as an initiative to change the culture of the mining sector, efforts to increase training of safety representatives, pressure from MHSI on mines to implement the outcomes of research undertaken through the MHSC, and the MHSI’s more forceful approach to enforcement (Vogt, et al., 2011; Creamer, 2012).

In addition, various companies had embarked on their own initiatives to improve health and safety (Faurie, 2011).

Without an established baseline and mechanisms to track improvements at site level, the impact of MOSH could not be ascertained reliably against this backdrop. The problem was identified in the evaluation study and persists, although other measures of MOSH progress have subsequently been embraced.

For example, the indicator for the number of adopting mines was poorly selected and the reach of the programme into the industry could not be assessed. The underlying reasoning for this was that individual mine sites were at different stages of adoption, some of which were complex and took months, if not years, to resolve, rendering the indicator ‘number of mines adopting’ meaningless over time. An exception was the leading practice for ‘entry examination and

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Figure 1 – Fatality rates per million hours worked, since agreement on the targets and milestones (adapted from DMR data, Msiza, 2013)

making safe’, which far outstripped other leading practices in terms of its spread into the industry. Notably, this practice necessitated a change in procedure rather than the introduction of new technology.

In addition to parallel initiatives and the absence of baseline data, the impact of MOSH was also obscured by changes in reporting requirements in the amended Mining Charter in 2010 (DMR, 2010), the subsequent 2012 Industry Mining Charter Health and Safety Report (Chamber of Mines, 2012), and an ’Adoption Scorecard’ introduced by the Adoption Teams within MOSH. Mines were required to report at first the number of mines adopting leading practice, and thereafter the proportion of adopting workplaces relative to the number of workplaces at which leading practices should be adopted.

However neither the Charter and nor the scorecard provided the necessary baseline to support meaningful measures of outcome and impact. The number of work sites adopting, number of teams trained, and units installed among others were the only measures of progress within MOSH. Appropriate outcome measures could be expected to include relevant health and safety leading indicators, and impact measures relevant lagging indicators of health and safety performance. However, none of this was possible without a baseline that enables tracking.

MOSH and the exercise of leadership within the health and safety regulatory framework The inception, implementation, and growth of MOSH should be understood within the context of health and safety regulation in the mining sector in South Africa.

South Africa’s health and safety legislative environment in the mining sector is guided by outcomes-based legislation that embodies general responsibilities for employers to provide safe and healthy workplaces, self-regulation, and participatory arrangements, allowing employers through engagement with workers to determine how particular outcomes are to be achieved. ‘A performance standard specifies the outcome required but leaves the concrete measures to achieve that outcome up to the discretion of the regulated entity.’ (Coglianese et al., 2002, p. 3) In other The Journal of The Southern African Institute of Mining and Metallurgy

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‘Progressive employers recognise that firms within an industry should work towards higher standards collectively, so that progress is not impeded by fear of unfair competitive advantage. There is virtually unlimited scope for practical work by industry-based safety bodies which can collate and interpret statistics, publish information, undertake technical surveys and research, and provide advisory services to individual firms, and liaise with government departments and inspectorates. Through these means each industry can work towards the solution of its own special problems.’ (Robens, 1972, p. 15) However, in the South African context, as MOSH sought to shift industry practices, it required the support and cooperation of labour and the state. Yet the initiative was conceived and promoted outside of the tripartite forums that enable government, labour, and the employers to work together. Both organized labour and the MHSI expressed a desire for MOSH strategy to be resolved through the MHSC, although this could change the voluntary status of the initiative. Nevertheless, as conceived in legislation, the MHSC’s functions are consistent with Robens Committee’s ideas.

The insistence of both labour and MHSI, that participation in MOSH should not be voluntary, was subsequent to the 2011 evaluation, and achieved by incorporating MOSH leading practices into the Mining Charter. The Charter is a regulatory instrument for ‘transforming’ the mining sector in South Africa. The scorecard and associated reporting template require companies to report on both the consideration and uptake of MOSH leading practices (DMR, n.d. (a), (b)). Thus the problem posed by the voluntary nature of MOSH of uptake at poorly performing mines was offset by developments in the regulatory environment.

The translation of a voluntary initiative into a regulatory requirement underscores a difficulty at the heart of the South African tripartite system, namely that the main employer body is not trusted (Shabangu, 2011) and has difficulty taking forward major voluntary initiatives without the expressed endorsement of stakeholders. Changes to the mandate of the MHSC, in which the role of the Minister of Mineral Resources (rather than the tripartite MHSC) is central to policy-making, reflect these dynamics (Odendaal, 2013).

In addition, stakeholders prefer an arms-length relationship with employers, as one way of being seen to preserve their own integrity. This has meant that MOSH is endorsed through enforcement rather than ongoing engagement over the content and impact of the programme.

Subsequent to the completion of the CSMI evaluation, the Learning Hub established, in addition to MOSH Task Team, the ‘MOSH Advisory Group’ to specifically include representatives from labour and the MHSI. However, this does not necessarily strengthen engagement. MOSH needs to deal with the problem of engagement and involvement of other stakeholders within the MHSC, rather than bypass it through establishing non-mandatory advisory structures.

Readiness for change, worker participation, and MOSH at the mine-site level The MOSH initiative reflects the Robens Committee’s assertion that ‘The primary responsibility for doing VOLUME 115

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words, it is not prescriptive. ‘In contrast to a design standard or a technology-based standard that specifies exactly how to achieve compliance, a performance standard sets a general goal and lets each regulated entity decide how to meet it’ (Coglianese et al., 2002, p. 3). Outcomes-based legislation creates space for national health and safety strategy. The 2013 health and safety milestones are an example of this.

South Africa is not alone in adopting an outcomes-based approach to health and safety legislation. The Mine Health and Safety Act (MHSA) 29 of 1996 is shaped by the approach to health and safety accredited to the British ‘Robens Committee’ (Robens, 1972) that concluded that ‘apathy’, both on the part of employers and employees, was significantly contributory to poor health and safety outcomes.

The recommendations of this Commission ushered in an era of health and safety practice that arises from the idea the there is a ‘natural identity of interests between employers and employees’ (Robens, 1972, p. 21) on issues of health and safety. The main features of the Robens model are that employers have primary responsibility for risk management, and that workers are party to decisions about how best to address health and safety risks through elected health and safety representatives and structures such as employeremployee health and safety committees. All of these are features incorporated into the MHSA. Although widely applied, the Robens model is also criticized as being too business-friendly, institutionalizing of the concept of selfregulation, and glossing over fundamental social conflict (Tombs and Whyte, 2012). The last point, about overlooking conflict, seems to have particular relevance for the South African mining sector, in which both class and racial divisions, often coinciding, play out.

To elaborate, South African outcomes-based health and safety legislation was introduced into a mining context shaped by apartheid and a skewed economic system that marginalized and divided people on racial grounds. Mining was at the heart of the apartheid economy, and miningrelated legislation until the late 1980s formally discriminated against black South Africans, who make up the majority of the population, by restricting them to manual work at the bottom of the employment hierarchy. Black mineworkers were confined to hostels and inferior rates of compensation were paid to black miners for occupational injury or illhealth. Estimates are that over 70 000 black miners died in mine accidents since the inception of mining, and an even larger number contracted silicosis. The 1994 Leon Commission of Inquiry into Safety and Health in the Mining Industry (Leon Commission, 1995) and the subsequent MHSA were early advancements associated with the postapartheid Mandela government. Sector-wide studies show that the apartheid legacy is still felt by workers (Hermanus et al., 2010; Shaw et al., 2010, 2011), especially with respect to racism, the practice of bonus payments, and a culture of blame. It is within this context that the South African experience of health and safety outcomes-based legislation is operative.

The MOSH initiative is in essence a voluntary industrylevel action, which the Robens Committee regarded as having high potential for improvements:

Mine Occupational Safety and Health Leading Practice Adoption System (MOSH) examined something about present levels of occupational accidents and disease lies with those which create the risks and those who work with them’ (Robens, 1972, p. 7), but runs into difficulties to those experienced elsewhere in respect of fully engaging the workforce. This is evident at both the national level of MOSH leadership (as discussed above) and at minesite implementation level as reflected in the findings. It is the latter that is discussed here.

One of the underlying assumptions of MOSH was that the MOSH process of engagement could induce a process of change. Numerous steps and templates in the MOSH handbook codified a process of engagement that encouraged practitioners to get in touch with what workers specifically were thinking and to ensure that mine-site leadership carried an appropriate message. However, findings in the evaluation point to the difficulty of achieving this without reference to the specific environment in which MOSH was being introduced. The contexts of each mine site varied, and hence the call by mines themselves for guidance about basic principles, and not ever-increasing detail. Therefore more flexibility was required which allowed mines to interface with the MOSH process differently, based on the level of readiness to embrace change.

Readiness is shaped by a number of factors (Hopkins, 2012; Novatis et al., 2012) It is shaped by whether managers at every level in an organization recognize the importance of dealing with health and safety risks and make this a priority; whether organizations have properly engaged workers to obtain their support and cooperation for efforts aimed at improving health and safety; and whether the experience and knowledge of workers of health and safety risks is understood and appreciated. It further extends to ensuring that processes are in place to secure and maintain the appropriate skills that are important in addressing health and safety risks. This also applies to different levels in an organization, so that workers, supervisors, managers, and technical support teams have complementary and integrated responsibilities for health and safety. Related factors that are equally important to an organization’s ability to address health and safety risks are quality of leadership, culture (especially a sense that organizational culture is just), trust, engagement, and worker participation (Gunningham, 2008; Dekker, 2008). These have specific relevance in South African mining, as discussed earlier.

Although MOSH embodied tools to address leadership practice, communication, worker perceptions, and attitudes, the problem of deep-rooted mistrust and all that flows from it were not specifically acknowledged or addressed. This could be achieved by linking to the industry-wide ‘Culture Transformation Framework’ (MHSC, 2011) which supports initiatives to shift the culture of the industry.

Linked to the issue of organizational readiness for change is the observation that the MOSH process did not incorporate the role of workers in health and safety as envisaged by the MHSA. At the site level, the expert-led paradigm of MOSH kept labour at the edge rather than central to improved health and safety. Worker health and safety representatives and employer-worker health and safety committees were not central to the MOSH process at the site level. Engagement with the site level health and safety committee and/or training of health and safety representatives specifically were not a feature of MOSH activities on site, but could be. During the evaluation, respondents from organized labour reported

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being brought into the MOSH process at a late stage. Meeting with the health and safety committee should be at an early stage of engagement and a key step in establishing buy-in at a mine.

MOSH collapsed the process of engagement with workers, to focus mainly on the collection of data to support the development of mental models about specific health and safety concerns and interventions. These models profile the knowledge, attitudes, and perceptions of workers and are used to shape behavioural communication and leadership behaviour plans. These types of studies can be timeconsuming and require that interviewers are properly trained to conduct interviews. One of the recommendations of the evaluation was that this type of study could be commissioned at a national/regional and/or commodity level rather than having to be repeated at each mine site. Communication and leadership plans could be adapted for specific mine sites.

However, more importantly this type of approach is fundamentally focused on building a process for change through understanding individual behaviour, when in fact health and safety is fundamentally about building an effective system (in which, obviously, consideration of individual behaviour has a role, but is not the driver).

Without sufficient integration into the existing health and safety system, MOSH effectively minimised the opportunity for engagement with labour. Engagement with workers should include active engagement with health and safety worker representatives and labour structures at the mine site, as well as with regional and national labour structures.

The extent to which the present MOSH initiative misread the process of engagement is further evidenced in the introduction to MOSH, subsequent to the 2011 evaluation, of ‘simple leading practices.’ Simple leading practices are examples of technology changes that can be made without needing to engage, such as winch covers to minimize dust.

This new focus was presented as ‘quick wins’, but also perpetuated a legacy of an industry unable to resolve issues of engagement and fundamental change, albeit in a difficult and complex environment.

Conclusions MOSH is an ambitious initiative with complex technical, structural, and procedural elements. It was clear the mining industry as a whole engaged with MOSH to various degrees, and labour and the MHSI saw MOSH as significant.

The full impact of MOSH and its sustainability across the mining sector was not clear. It was evident that where MOSH worked well, mines usually considered and implemented several leading practices simultaneously. This was particularly the case when a mine subscribed to a broad strategy of improvement and culture change. In these instances there appeared to be greater capacity to absorb the processes and technologies of the MOSH adoption system. Difficulties in assessing the impact of MOSH arose from many quarters within the MOSH programme itself, which did not clearly define and measure its impact, and the fact that MOSH was one of many initiatives to improve health and safety in South African mines. Many of these were launched and operated simultaneously, and the effect of one on the other was never considered. An example of this was how a tough new regulatory approach, the culture transformation framework, and the MOSH processes could be expected to interact.

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References CHAMBER OF MINES. Not dated. MOSH - Learning From Each Other by Learning Practice Adoption. http://www.mosh.co.za/ [Accessed 13 January 2014].

CHAMBER OF MINES. 2012. Mining Charter, 2012 Health and Safety Report.

Chamber of Mines of South Africa, Johannesburg.

COGLIANESE, C. and LAZER, D. 2003. Management-based regulation: prescribing private management to achieve public goals. Law and Society Review, vol.

37, no. 4. pp. 691–730.

COGLIANESE, C., NASH, J., and OLMSTEAD, T. 2002. Performance-Based Regulation.

Prospects and Limitations in Health, Safety, and Environmental Protection. Regulatory Policy Report no. RPP-03. Harvard Kennedy School, Harvard University, Cambridge MA. CREAMER, M. 2012. Section 54 stoppage issue dominates Shabangu media conference. Mining Weekly, 20 March 2012. www.miningweekly.com [Accessed 17 January 2014].

CREAMER, M. 2013. SA mines Installing 170 000 safety bolts a day, 200 km safety netting. Mining Weekly, 27 November 2013. www.miningweekly.com [Accessed 10 December 2013].

DEKKER, S. 2008. Just Culture. Who gets to Draw the Line? Springer-Verlag, London.

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DMR (Department of Mineral Resources). 2010. Amendment of the Broadbased Socio-Economic Empowerment Charter for the South African Mining and Minerals Industry. Pretoria.

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Mine Health and Safety Inspectorate, Pretoria.

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DMR (Department of Mineral Resources). Not dated (a). Mining Charter Reporting Template, Final. Pretoria.

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GUNNINGHAM, N. 2008. Occupational health and safety. Worker participation and the mining industry in a changing world of work. Economic and Industrial Democracy, vol. 29, p. 336.

HERMANUS, M., COULSON, N., and MAGNER, C. 2010. Creating a Shared Approach to Improving Safety in Mining “A White Flag Every Day”. CSMI, University of the Witwatersrand and REOS, Johannesburg HOPKINS, A. 2012. Disastrous Decisions -The Human and Organisational Causes of the Gulf of Mexico Blowout. CCH, Australia.

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MINE HEALTH AND SAFETY COUNCIL. 2011. Zero Harm through Action - Culture Transformation Framework for the South African Mining Sector.

Johannesburg.

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NOVATIS, E., MCCORMICK, P., WOODSIDE, E., and LARDNER, R. 2012. Beyond Bulletins and Presentations: Use of Scenarios to Learn from Accidents.

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PIENAAR, G. J. and DU PLESSIS, J.J. 2009. Meeting and exceeding the occupational hygiene milestones. Hard Rock Safe Safety Conference 2009. Southern African Institute of Mining and Metallurgy, Johannesburg. pp. 231–234.

ROBENS, L. 1972. Safety at Work. Report of the Committee. Government Report, United Kingdom. Health and Safety Executive, Bootle, Merseyside, UK. SHABANGU, S. 2011. Keynote Address by Ms S Shabangu, Minister of Mineral Resources, at the Mine Health and Safety Council (MHSC) Summit at Emperor's Palace, 17 November 2011.

SHAW, A., BLEWETT, V., and SCHUTTE, S. 2011. International Frameworks, National Problems: Mining OHS regulation in South Africa and Australia.

34th International Conference of Safety in Mines Research Institutes, New Delhi, India, 7-10 December 2011. pp. 545–554.

SHAW, A., SCHUTTE, S., COX, S., BLEWETT, V., MILANZI, L., MORABA, T., FORMANOWICZ, A., and MOKOENA, A. 2010. Changing Minds, Changing Mines: Final Report to the Mine Health and Safety Council. Mine Health and Safety Council, Johannesburg.

TOMBS, S. and WHYTE, D. 2012. Reshaping health and safety enforcement: better regulation. Making Employment Rights Effective. Dickens, L. (ed.). Hart Publishing, Oxford, UK. pp. 67–86.

VAN DER WOUDE, S. 2014. Employer Summit 2006, Improvment Required for Milestones.

VAN DER WOUDE, S., STEWART, J., ALLY, I., and LEBURU, T. 2004. Strategies for Sustainable Improvement in Safety Performance. 0141206-COM-Safety paper-final 6 Dec 04- SvdW-c7. Chamber of Mines, Johannesburg.

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There are also factors that hamper MOSH which are specific to the operational environment of South African mines and the nature of the Chamber of Mines. ➤ Although the Chamber represents the majority of mine employers, it is unable on its own to take forward national-scale decisions even when these may be in the best interests of the workers and people affected by mining. Thus one of the unspoken challenges associated with MOSH at the time of the evaluation was that it was championed by an employer body with primarily an advocacy role, which relied solely on voluntary uptake by mining companies ➤ As found in the evaluation, the Chamber, as a voluntary organization representing employers, could not set hard performance targets for mines.

Consequently the costs and resources necessary to sustain MOSH over time were likely to have diminishing returns because the initiative could not target companies with the poorest health and safety records, which in turn could chose not to adopt MOSH for their own reasons ➤ While the MHSI chose not to participate in the design and planning of MOSH, or actively lead the uptake of leading practices, two years after MOSH’s establishment it provided indirect support through regulation. The incorporation of MOSH into the Mining Charter and related scorecard introduced pressures to comply, something that the Chamber was unable to do.

The effect of this might still be, ironically, to mask and diminish the problem of not being able to target mines where performance is wanting.

Another factor that affects MOSH was that the programme did not formally recognize that different organizations were not able to initiate and absorb change in the same way. However, in the Adoption Teams and on mine sites, the organizational context of health and safety did come into play, and found expression in the concerns raised over the complexity and highly-structured nature of the process, and the adaption of the process on site.

Finally, it is apparent that the MOSH process unfolded outside the formal health and safety system described in the MHSA. Whether mines can find ways of drawing health and safety committees and health and safety representatives into the programme is a matter worthy of further consideration.

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http://dx.doi.org/10.17159/2411-9717/2015/v115n8a9

Application of the attainable region technique to the analysis of a full-scale mill in open circuit

by F.K. Mulenga* and M.M. Bwalya†

The application of the attainable region (AR) technique to the analysis of ball milling is currently limited to batch data. This paper introduces the use of the technique to continuous milling.

To this end, an industrial open milling circuit processing a platinum ore was surveyed. Samples were collected and later characterized by means of laboratory batch testing. On site, several milling parameters were varied systematically so as to collect data for modelling purposes. These paramters included ball filling, slurry concentration, and feed flow rate.

After data analysis, a simulation model of the open milling circuit was developed under MODSIM®, a modular simulator for mineral processing operations. The mill was then simulated and the data generated was analysed within the AR framework.

Initial findings reveal an opportunity to gain valuable insight by studying milling using the AR technique. From an exploratory perspective and inasmuch as this study is concerned, feed flow rate, ball size, and ball filling were identified as being pivotal for the optimization of open ballmilling circuits. Mill speed, on the other hand, had only a limited effect on the production of particles in the size range -75 +10 μm.

Keywords attainable region, ball milling, population balance model, milling parameters, scale-up procedure, MODSIM® simulator.

Introduction Glasser and Hildebrandt (1997) propounded the attainable region (AR) as a technique for the analysis of chemical engineering reactor systems. Results have since been produced and tested on both the laboratory and pilot scales. The method results in a graphical description of chemical reactions by considering the fundamental processes taking place in the system, rather than the equipment.

From the plotted graphs, the process and the reactors can be synthesized optimally into a flow sheet.

The use of the AR method still has a long way to go as far as mineral processing is concerned. For instance, several articles have reported the application of the AR technique to ball milling: Khumalo et al. (2006, 2007, 2008); Khumalo (2007); Metzger et al. (2009, 2012); Metzger (2011); Katubilwa et al.

(2011); Hlabangana et al. (2012). Their main shortcoming has been the exclusive use of laboratory batch grinding data.

To address this deficiency, Mulenga The Journal of The Southern African Institute of Mining and Metallurgy

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* Department of Electrical and Mining Engineering, University of South Africa. † School of Chemical and Metallurgical Engineering, University of Witwatersrand, South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received June 2014 and revised paper received March 2015.

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Synopsis

and Chimwani (2013) proposed a way by which the technique could be extended to continuous milling. In effect, the batch milling characteristics of a platinum-bearing ore (Chimwani et al., 2013) were used and scaled up to an open milling circuit. Then, with simplifying assumptions, an attempt was made to optimize the residence time of particles inside the mill. Later, Chimwani et al. (2014a) presented some optimization examples involving various milling parameters. The sequence of published articles then paved the way for the study of industrial milling systems with the AR methodology.

Admittedly, the limitation has been that the exit classification of the milling circuit was not included (Mulenga and Chimwani, 2014; Chimwani et al., 2014a, 2004b). The importance of this internal phenomenon has been discussed in detail elsewhere (Cho and Austin, 2004; Austin et al., 2007). Suffice to say that the exit classification (also referred to as post-classification or internal classification) is responsible for the preferential discharge of smaller particles and the retention of larger particles back into the mill load until sufficient milling has been achieved.

In the present work, MODSIM® – a modular software package for the simulation of mineral processing units (King, 2001) – was used. The flexibility of the simulator enabled the internal classification of particles before exiting the mill to be taken into account, thereby making it possible to generate industrially sound data. From there, the effects of ball filling, ball size, mill speed, and feed flow rate on the product of an open milling circuit were simulated.A methodology for the

Application of the attainable region technique to the analysis of a full-scale mill in open circuit application of the AR technique to realistic open mill models is presented and illustrated using the simulated data. The AR profiles produced are analysed and interpreted within the AR framework. As a result of this analysis strategy, new insights on open milling circuits are brought to light and discussed.

Finally, recommendations for future work are formulated.

Theoretical background This section reviews the theoretical modelling of batch milling as well as the fundamental processes associated with milling in general. The basic concepts underpinning the AR technique are also discussed.

Ball milling model The objective of any comminution operation is to break large particles down to the required size. In tumbling ball mills, this is achieved through repetitive breakage actions.

Fragments from each particle generated after the initial breakage actions generally fall into a wide range of sizes.

However, some of the daughter fragments are still coarse and require further breakage. That is why the milling process can be regarded as the combination of two simultaneous actions: the selection of particles for breakage, and the actual breakage resulting in a particular distribution of fragment sizes after the particle has been selected (Gupta and Yan, 2006). A size-mass balance inside the mill that takes into account the two aforementioned reactions eventually results in the full description of the grinding process.

Let us consider particles of size xi at time t within a given feed size distribution; denote their mass fraction wi(t). Now, take a time interval dt small enough to allow only single breakage events to occur on a fraction of wi(t).

If the fraction selected for single event breakage per unit time is Si, then Si.dt represents the mass fraction broken after the time interval dt. This mass breaks into a wide range of child particles, the size of which spans from the parent size xi down to xn.

In an open ball-milling circuit, the flow of material in and out of class interval [xi, xi+1] can be divided into four categories: the mass fraction accumulated because not selected for breakage, the mass fraction leaving the class interval as a result of the single breakage events, the incoming mass fraction through breakage of particles larger than xi, and the incoming mass fraction from the new feed.

After time interval dt, the second category of particles is given by Si.dt, while the third category necessitates the determination of the mass fraction reporting to [xi, xi+1] as a result of the breakage of selected particles of initial sizes larger than xi.

The term Si is the selection function; it represents the rate of breakage of particles of size xi. Austin et al. (1984) proposed the following empirical model to express the variation of the selection function with particle size:

[1] where xi is the upper size of the particle size interval i under consideration a and μ are parameters that are mainly functions of milling conditions α and Λ are material-dependent parameters.

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Let us now assume that xj represents any particle size larger than xi (that is, xn ≤ xi ≤ xj). It is clear that the particles of size xj selected for breakage will give birth to particles of sizes spanning from xj down to xn.

Because particle size is conventionally measured using a series of sieves with mesh apertures arranged in a geometric sequence (generally 21/2 or 21/4), a convenient notation will be introduced whereby the largest size class interval is named x1. Particles in this class pass through a sieve of size x1 but are retained on a sieve of size x2. Consequently, the mass fraction of particles falling in size class interval [x1, x2], or in class 1 for short, at time t becomes w1(t). The last size class interval, known as the ‘sink’ and composed of the smallest particles, is termed [xn, 0] or class n. Particles in the sink class are therefore of size xn and their corresponding mass fraction is wn(t).

In order to define the breakage function, consider two class intervals [xi, xi+1] and [xj, xj+1] containing particles of size xi and xj respectively where xi < xj. The breakage function, better called the primary breakage distribution function, can be defined as the average size distribution resulting from the fracture of a single particle (Kelly and Spottiswood, 1990). It is used to describe the size distribution of the child particles produced after a single step of breakage of a parent particle of the material under consideration. Hence, if a parent particle is impacted by a grinding ball, the resulting product will consist of broken particles in a wide size range. The description of this breakage event (single step of breakage) is made possible by defining the breakage function of the material being broken. To this end, the primary breakage distribution function of particles of size xj breaking into size xi is defined as follows: [2] A more convenient way of describing the breakage distribution function is to use the cumulative breakage function, defined as follows (Austin et al., 1984): [3] With this new definition, the following empirical model relating the cumulative breakage function to particle size can be used (Austin et al., 1984): [4] where β is a parameter characteristic of the material used γ is also a material-dependent characteristic Φ represents the fraction of fines produced in a single fracture event. It is also dependent on the material used.

Attainable region analysis applied to ball milling The AR methodology was intended primarily for chemical process optimization. That is why its original utilization entailed the concentrations of reactants and products of interest. In milling, however, particles break into different size classes. Hence, by analogy between chemical reactions and milling, size classes can be regarded as chemical species.

In doing so, the feed size class becomes the reactant and the The Journal of The Southern African Institute of Mining and Metallurgy

Application of the attainable region technique to the analysis of a full-scale mill in open circuit

Figure 1 – Particle size distribution of a silica sand tested in the laboratory (data from Khumalo, 2007)

Figure 2 – Grinding kinetics as plotted for the three size classes m1, m2, and m3 The Journal of The Southern African Institute of Mining and Metallurgy

Let us consider a given passing sieve size xk, where xk ∈ {xi with 1 ≤ i ≤ n}, and define new size classes: ➤ The feed size class or the material of size x falling between x1 and x2; the mass fraction of material in this class will be termed m1 ➤ The middling size class constituted of particles of size x between x2 and xk; the corresponding mass fraction will be termed m2 ➤ The fines size class m3, which is defined as the mass fraction of material passing through screen size xk.

For illustration purposes, let x2 and xk be 4000 μm and 600 μm respectively. The change in mass fraction in each new size class (i.e. m1, m2, and m3) can now be tracked as a function of grinding time: this will look something like Figure 2.

If the objective is, say, to produce as much material in the middling class as possible, grinding the feed for 5 minutes would be the way to go, as shown in Figure 2. Note that a model is not required to find such an optimum; instead, a straightforward exercise of interpreting and reading graphs does the work.

From the mass fraction profiles reported in Figure 2, let us present the data in the last format, which is of much interest to the AR technique. In order to perform the transformation, respective mass fractions are read off from Figure 2 at grinding time t = 30 minutes. They are then mapped onto a two-dimensional mass fraction space, that is, (m1, m2). By repeating the process and mapping all the data points in Figure 2, the AR plot is produced as shown in Figure 3. This figure presents the same data produced from the silica sand material in Figure 2, but this time in the two-dimensional space (m1, m2).

It should be recalled that m3 can be inferred by mass balance at any stage of the process. Most importantly, Figure 3 represents the path followed by the milling process to achieve the specified objective function from the system feed. This is referred to as the ‘attainable region path’.

The star marker on the plot (in Figure 3) corresponds to the 30 minutes’ grinding time considered earlier. From an AR perspective, one would say that about 97% of the feed is needed to produce approximately 13% of middling; and this, after 30 minutes of grinding. It is therefore understood that after 30 minutes only 3% mass fraction is left in the feed class m1 while 13% of middling m2 is produced.

Figure 3 – AR plot relative to the same silica sand tested with xk = 600 μm VOLUME 115

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undersized classes are the products of the breakage reaction.

Glasser and Hildebrandt (1997) define the attainable region as ‘the set of all physically realizable outcomes using only the processes of reaction and mixing in steady-state systems for some given feed(s)’. In other words, given the feed and the reaction kinetics, the set of all possible outputs of a chemical reaction can be determined; and from there, the best operating conditions (subject to some external constraints) can be deduced.

In order to illustrate the paradigm, let us start off with a narrow-sized feed material [x1 ≤ x ≤ x2]. If after batchmilling the feed sample for a grind time t = 1 minute, a complete particle size analysis is performed from x1 down to xn, Figure 1 can be plotted.

Figure 1 does not only present the product size distribution (PSD) for t = 1 minute, but also shows the PSDs corresponding to grinding times ranging from 2.5 to 40 minutes for which particle size analyses were performed in a similar fashion.

Application of the attainable region technique to the analysis of a full-scale mill in open circuit If the objective is to produce at least 40% of m2, the AR plot indicates that between 44% and 89% of the feed material is required. The graph shows that, for the objective to be met, between 11% and 56% of the material should remain in class m1. This translates to a mass fraction of between 89% and 44% of material that needs to be ground out. Nonetheless, information such as the specific energy to be used, or some other operating constraints, will orientate the process engineer towards the choice of the right mass fraction to be milled.

It appears that AR paths make it possible to characterize the selectivity of the process. In other words, it is possible to determine the fraction of initial feed material that will report to the size class of interest under given operating conditions.

It is important to note that the AR analysis presented here is for illustration purposes, and that each case study will require particular attention.

The last point to discuss is the idea behind the term ‘attainable region’. This can be conceived as follows. In Figure 4 (created from Figure 2), point A represents a fresh feed not yet ground, while point B represents a product milled for some time that consists of 15% m1 and 50% m2. It is possible to mix a fraction of A, say 25%, and combine it with 75% of B to obtain a composite material C. One can carry on with this exercise and fill up the region between the AR plot and the x-axis with data points for different combinations.

The data points will represent all the possible composite materials that can be obtained from the milling system using the mixing principle. That is why the stripped region in Figure 4 is called the attainable region.

If the AR region is concave (note that the striped region in Figure 4 is convex), mixing presents an advantage in that more material can be produced in addition to that generated by the grinding process itself. Furthermore, the maximum point M (in Figure 4) can be achieved only if the class of interest is defined with upper and lower screen size boundaries. As a corollary to this, the sink fraction (i.e. class n) and the feed fraction (i.e. class 1) never experience such a maximum. That is the reason why a definite class interval is always required for process optimization, and not a semiinfinite class. Sink and feed fractions are typical examples of semi-infinite classes.

Data collection methodology This article aims to present a way of applying the AR technique to an open milling circuit. To this end, a full-scale mill, as well as the ore being processed, were characterized.

The data collection technique is discussed in the subsequent sections, together with the technological specifications of different experimental set-ups. The simulation package employed in the modelling of the full-scale mill is also presented succinctly. Finally, the input parameters for the actual ball mill model are listed.

Full-scale milling data The full-scale mill considered in this work is an overflow discharge mill, run in open circuit and used in the secondary milling of a UG2 platinum ore. The technological and operating specifications are: Mill rated power 11 000 kW Mill full length L = 9.6 m inside liners Mill diameter D = 7.312 m inside liners Mill speed φc = 75% of critical Ball filling J ranging from 25% to 33% Steel ball diameter d = 40 mm.

The mill is lined with 44 rubber lifters that have a height of 100 mm. The solids concentration in slurry is on average 75% by mass for a solids feed rate, F, of 330 t/h.

The set of industrial parameters that were monitored is shown in Figure 5, together with the flow sheet of the secondary milling section. Data on the density, the flow rate, and the size distribution of the densifier underflow stream was collected at the mill inlet, as well as the flow rate of the mill dilution water. At the mill discharge, the density, flow rate, and size distribution of the mill product were measured.

In this work, the aforementioned industrial data was used to validate the simulation model of the open milling circuit in Figure 5. It is important to mention that the full-scale milling data collected was the result of a fruitful collaboration between Magotteaux (Pty) Ltd, the University of Cape Town (UCT), Anglo Platinum’s Waterval UG2 Concentrator, and the University of the Witwatersrand (Wits). More details on the industrial campaign and the sampling set-up are provided in Keshav et al. (2011).

Batch milling data The complete description of milling necessitates an accurate measurement of all parameters involved in the process model.

A good way of doing this is to run well-planned batch tests on mono-sized feed samples using the one-size-fraction method (Austin et al., 1984). In this method, a sample in one size class is prepared and loaded in a small laboratory mill together with grinding balls. Milling is performed for several suitable grinding time intervals. After each interval, the product is sieved, returned to the mill, and then further milled. In this way, the mill product is monitored for the different grinding time intervals chosen a priori. Lastly, laboratory results are analysed in order for the selection function parameters (i.e. aT, μT, α, and Λ in Equation [1]), as well as the breakage function parameters (i.e. β, γ and Φ in Equation [4]), to be determined. That is why the primary objective of batch grinding tests is to measure the milling characteristics of a material under given experimental conditions.

Figure 4 – Principle of mixing in the AR space

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Application of the attainable region technique to the analysis of a full-scale mill in open circuit

Figure 5 – Set-up of the full-scale mill, sampling points, and type of data collected (Keshav et al., 2011)

Table I

Milling parameters of the platinum ore used (Chimwani et al., 2013) Breakage function parameters

Selection function parameters

β γ Φ aT μT α Λ

6.2 1.50 0.60 0.42 1.12 1.37 4.74

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information was used in setting up the simulation model for the open milling circuit in Figure 5.

If one considers a ball mill of internal volume Vmill, it is clear that the mill can theoretically carry an equal volume of grinding media to its own volume. In practice, however, only a fraction of the volume, Vballs, is occupied by grinding balls.

The ratio of the volume occupied by balls at rest to the mill volume is defined as the ball filling, J. In addition to the bed of balls, and specifically in wet milling, slurry (which is a mixture of ore particles and water) is also loaded into the mill. Depending on the volume loaded, slurry occupies firstly the interstices between the grinding balls before immersing the bed of balls at rest. The ratio of the volume of slurry loaded to the volume of ball interstices available within the bed at rest is the slurry filling, U. Hence, considering a ball filling J, the volume of slurry Vsl needed to achieve slurry filling U is determined as follows: [5] where ε represents the porosity of the bed of grinding media at rest, which is assumed to be 0.4 on average.

A limitation to the gradual increase in ball filling is that the maximum power occurs when the media filling, J, is at approximately 45% of the mill volume; after that, power tends to decrease. By the same token, slurry should preferably be loaded in a way that ensures that most of the material is held in the media interstices (Latchireddi and Morrell, 2003). Otherwise, if the level of slurry is low, the charge is more likely to experience ball-to-ball contact and waste energy. On the other hand, if there is more slurry than the media charge can hold within, a pool of slurry will form, with impact breakage becoming less pronounced.

Another important milling parameter that needs a brief explanation is the rotational speed. Mill speed is commonly defined as a fraction of the theoretical critical speed of the mill. The critical speed is the speed of the mill at which a single ball starts to centrifuge. At this speed, the grinding ball VOLUME 115

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In the present work, the material used in the batch testing programme was the platinum-bearing UG2 ore already subjected to primary milling. The UG2 is one of the platinumrich layers in the Bushveld Complex (BC) of South Africa and accounts for about 60% of South Africa’s platinum reserves.

Internal references from the Waterval concentrator reported that the specific density of the UG2 ore used is 3.47 kg/cm3 as established in routine inspection. The ore supplied was a fine material of size less than 1700 μm.

Batch testing was conducted on three mono-sized ore samples: -850 +600 μm, -600 +425 μm, and -425 +300 μm.

Three ball diameters were used: 10 mm, 20 mm, and 30 mm.

The feed samples were dry batch-milled for 0.5, 1, 2, 4, 8, 15, and 30 minutes. After each grinding time, a representative sample was taken from the mill powder for conventional sieving. In total, nine series of batch tests were carried out for seven grinding times.

Table I lists the breakage function and selection function parameters characterizing the UG2 ore used. Chimwani et al.

(2013) discussed in detail how these parameters were determined.

The values in Table I were measured under the following experimental conditions: ball diameter dT = 20 mm; ball filling JT = 20%; powder filling UT = 0.75; mill speed φcT = 75% of critical; and mill diameter DT = 302 mm. All this

Application of the attainable region technique to the analysis of a full-scale mill in open circuit sticks against the mill wall because the centrifugal and the gravitational forces are in balance. Critical speed is given in revolutions per second by the following expression: [6] where D is the internal diameter of the mill d is the diameter of the grinding ball under consideration g is the gravity constant, i.e. 9.81 m/s2.

Therefore, if the mill is said to run at 70% of critical, this simply means that the actual speed of the mill, expressed in revolutions per second, is 70% of the theoretical critical speed calculated using Equation [6].

MODSIM®, the modular simulator The core of the investigative work was centred on the use of the academic version 3.6 of MODSIM®, a specialized steadystate simulator. The software package is a modular simulator for ore-dressing plants. King (2001) provides a comprehensive review of all the unit processes available in MODSIM® as well as a description of relevant models applicable to each operation.

As far as the milling circuit model is concerned, GMSU was found to be the adequate option for this exploratory study. The

MODSIM® ball mill model is an encapsulation of the scale-up procedure by Austin et al. (1984). It is used when the selection function and breakage function parameters have been determined from laboratory batch tests. In addition to this, the dimensions of the full-scale mill should be available. This is indeed the case in the present work. Most importantly, the GMSU model assumes that post-classification is present and that the mill load is perfectly mixed.

Figure 6 shows the form for the input of parameter values necessary to the GMSU model. In this study, no liberation model was considered; furthermore, provision was made for mill overfilling.

In the execution of all simulations, unless otherwise stated, the standard ball size distribution available in MODSIM® and the feed size distribution in Figure 7 were used.

Figure 7 represents an average feed size distribution calculated from the various feed samples collected on the plant during the sampling campaign described previously.

Simulation results Simulation model validation This section presents the development of the MODSIM® ball milling model as well as aspects of the model validation undertaken. The effects of selected milling conditions are discussed in subsequent sections.

Figure 6 – Setting up the MODSIM® model of the mill

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Application of the attainable region technique to the analysis of a full-scale mill in open circuit

Figure 7 – Feed size distribution used for simulation purposes

As a starting point, a simulation program for the open milling circuit in Figure 5 was initiated using MODSIM®. To this end, the breakage characteristics of the ore (see Table I) were declared for scale-up as shown in Figure 6. The product size distributions were then generated for different operating conditions.

Figure 8 illustrates a typical simulation output window rendered by MODSIM®.

Thereafter, the AR methodology was applied to the data with the objective of assessing the influence of several milling conditions on the production of material amenable to flotation from the initial feed size distribution shown in Figure 7. The product class m2 was set between 75 μm and 9 μm because the platinum industry in South Africa generally requires the product to be below 75 μm before it is sent to flotation. The cut-off size of 9 μm was guided by the poor flotation performance reported for particles less than 10 μm on average (Rule and Anyimadu, 2007).

All in all, the initial feed class m1 considered was -1700 +75 μm, while the objective function was to explore the production of m2, i.e. -75 +9 μm, as a function of ball filling, feed flow rate, mill speed, and ball diameter.

Before reporting the findings pertaining to the AR method, simulated results from the MODSIM® milling model were compared to the industrial data. It can be seen in Figure 9 that model and measurement agree well for particle sizes larger than 100 μm. In contrast, below this size, measurements are under-predicted by the simulation model.

The model predicts a coarser mill product compared to the industrial data.

This is also evident in Table II, in which the predicted 50% passing sizes (d50) are higher than the experimentally measured ones. Possible reasons for these discrepancies are presented later in the Discussion section. However, agreement on the 80% passing size (d80) is deemed enough (see Table II) to carry out exploratory simulations and meet the objectives set for the present article.

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Figure 8 – Example of PSD output rendered by MODSIM® under the following simulation conditions: Cw = 75.1%, F = 205.1 t/h, J = 29.9%, and φc = 75% of critical speed

Application of the attainable region technique to the analysis of a full-scale mill in open circuit

Figure 10 – Effects of ball filling on mill product under the following simulation conditions: F = 330 t/h, Cw = 70%, φc = 75% of critical speed, and standard ball size distribution

Figure 9 – Simulated and experimental mill product under the following conditions: Cw = 75.1%, F = 345.5 t/h, J = 29.9%, and φc = 75% of critical speed

Effects of ball filling on mill throughput In the first set of simulations, the effect of ball filling on the mill throughput was assessed. Four levels of ball filling were simulated: J = 10, 20, 30, and 40%. The term ‘throughput’ is used to refer solely to the mass fraction of particles in class m2 present in the mill product. It is an indication of the ability of the mill to produce the desired particles, i.e. -75 +9 μm. It does not consider the mass or volume flow rate of the mill discharge stream.

It can be seen in Figure 10 that the mill product size distribution becomes finer as ball filling is increased. Note here that feed flow rate, ball size distribution, slurry filling, and mill speed were kept constant.

Now, examination of the data in Figure 10 from an AR point of view reveals that the production of m2 follows a straight line which is close to the ideal AR profile. In effect, the ideal AR profile is represented in Figure 11 by the red dotted line: along this line, the feed is milled in such a way that the product reports to class m2 only. In other words, there is no loss to finer sizes or no material is produced below 10 μm.

Figure 11 – Attainable region profile showing the effects of ball filling on the mill product. Simulation conditions: F = 330 t/h, Cw = 70%, φc = 75% of critical speed, and standard ball size distribution

Table II

Measured and simulated size characteristics of the mill product Milling parameters

d80 [μm]

d50 [μm]

J [%]

F [t/h]

Cw [%]

Measured

Modelled

Measured

Modelled

24.5

206.9

75.6

106

119

48

85

24.7

199.8

71.4

115

117

58

83

24.6

211.2

74.5

110

115

52

83

24.9

191.1

67.3

111

115

57

81

29.1

180.0

73.0

105

101

44

66

29.5

196.5

65.4

110

102

48

68

29.9

205.1

75.1

107

103

48

69

30.2

203.9

73.4

105

102

47

68

30.2

222.1

65.0

110

104

52

70

30.2

207.2

63.5

110

102

52

68

32.9

199.4

67.7

111

98

52

63

32.8

209.8

65.1

110

99

54

65

32.9

191.1

75.6

100

96

43

62

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Application of the attainable region technique to the analysis of a full-scale mill in open circuit Because the AR plot in Figure 11 is close to the ideal AR plot, it can be argued that the effect of ball filling as far as the platinum ore and the full-scale mill set-up are concerned is such that particles are preferentially broken into m2. Three data points are also labelled to assist with the graphical interpretation.

Note that the identification of the optimum ball filling is not straightforward from Figure 11. However, it becomes easy to see the optimum once the data is presented using ball filling as the independent variable (see Figure 12).

In addition to the above, the simulation results suggest that the optimum ball filling is somewhere between J = 35% and J = 40% as shown in Figure 12. In this range, the production of m2 is as high as 56.5%.

Effects of mill speed on mill throughput The next series of simulations was aimed at investigating the effect of mill speed on the production of m2. The mill speed values considered spanned from 50% to 90% of critical speed.

Similarly to the effect of ball filling examined previously, the AR profile in Figure 14 indicates a limited effect of mill speed on the throughput. However, presenting the production of m2 as a function of mill speed proves to be insightful.

Indeed, Figure 15 shows an increase in m2 with mill speed until 80% of critical speed; then a steady drop follows.

Effects of feed flow rate on mill throughput

Figure 13 – Attainable region profile showing the effects of solids feed rate, F, on the mill product. Simulation conditions: Cw = 70%, J = 30%, φc = 75% of critical speed, and standard ball size distribution

Figure 12 – Effects of ball filling on the production of material in class m2. Simulation conditions: F = 330 t/h, Cw = 7 %, φc = 75% of critical speed, and standard ball size distribution

Figure 14 – Attainable region profile showing the effects of mill speed on the mill product. Simulation conditions: F = 330 t/h, Cw = 70%, J = 30%, and standard ball size distribution

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In the second set of simulations, the effect of solids feed rate was analysed while keeping ball filling, slurry filling, and mill speed constant.

For the sake of exploring the relevance of the AR technique, the flow rate was varied between 10 t/h and 400 t/h. It can be seen in Figure 13 that although the full-scale mill has been operated at a flow rate of the order of 300–400 t/h, better throughput may be obtained at flow rates as low as 30 t/h. Furthermore, the mass fraction of m2 in the mill product increases from an average of 40% to in excess of 75%. This is indicative of the fact that lower flow rates are conducive to the production of -75 +10 μm particles.

Vermeulen et al. (1991) reported similar findings using mills of different sizes in open circuit. The only difference is that the target product size was less than 75 μm. The present work, on the other hand, focuses on particles between 75 μm and 10 μm.

Most importantly, even though a better mill product is obtained at low flow rates, the volume produced may be too low to justify the implementation of the AR optimum. A study is currently underway to find a reasonable trade-off.

Application of the attainable region technique to the analysis of a full-scale mill in open circuit

Figure 15 – Effects of mill fractional speed on the production of material in class m2. Simulation conditions: F = 330 t/h, Cw = 70%, J = 30%, and standard ball size distribution

Figure 16 – Attainable region profile showing the effects of ball size on the mill product. Simulation conditions: F = 330 t/h, Cw = 75%, J = 30%, and φc = 75% of critical speed

Figure 15 also suggests that in order to obtain maximum production of m2, mill speed should be adjusted to about 80% critical. The optimum operating range as far as m2 is concerned is therefore 75–85% of critical speed.

(d50) was over-predicted. One possible reason for the systematic discrepancies observed is believed to be related to the modelling of the internal classification. Indeed, the general consensus is that the exit classification is predominantly a function of particle size (Austin et al., 1984; King, 2001; Napier-Munn et al., 1996; Cho and Austin, 2004; Cho et al., 2013). One of the widely used models for the exit classification is the logistic function given in Equation [7] (Austin et al., 1984):

Effects of ball size on mill throughput The last of series of simulations sought to ascertain the effect of ball diameter on mill throughput. To this end, the diameter of grinding balls was varied between 10 mm and 50 mm.

It can be seen in Figure 16 that smaller grinding balls encourage the production of m2 while larger ones basically break particles indiscriminately. The problem with smaller balls is that their life span inside the mill is short, and therefore, makes their use less recommendable. That is why it is believed that a ball mix with a high proportion of small balls would be a viable option (Cho et al., 2013). The mixture of balls will take advantage of the number of small balls while extending their life with a fair amount of larger ones.

Discussion Attempts to apply the AR technique to comminution have been undertaken in the past with interesting outcomes. In particular, a viable approach has been proposed (Mulenga and Chimwani, 2013; Chimwani et al., 2014a, 2014b) and is being further developed. The shortcoming of this series of articles has been the exclusion of the internal classification at the mill exit.

Unlike the previous papers, the present work has utilized a software package that is able to simulate open milling circuits while allowing for exit classification. It has thus been possible to generate sound industrial data for analysis within the AR framework. Nonetheless, the simulation needed validation before use. The outcomes are shown in Figure 9 and Table II. As an exploratory study, it is argued that the simulation model is able to predict satisfactorily the 80% passing size (d80) of the mill product. However, the 50% passing size

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[7] where xi is the upper size of the particle size interval i under consideration d50 represents the cut-off size of the post-classification. It can be regarded as the size at which particles have equal chance of reporting to the mill product or back into the mill load. Thus, c(d50) = 0.5 λ is a parameter defining the gradient of the classification function c(xi) when plotted as a function of particle size xi. The gradient is related to the sharpness index S.I. as follows:

In the MODSIM® model of the mill, the following classification parameters were found to produce good results: d50 = 128.6 μm and S.I. = 0.35.

To return to the discrepancies in Figure 9, one can see that Equation [7] is intrinsically dependent on particle size xi.

It is possible that the actual exit classification of the mill could be affected by operating parameters other than particle size. If that is the case, then the classification model should be revised and the predicted mill product adjusted. Work is currently being conducted in which the classification function 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

Conclusion and future outlook The main objective of the present work was the proposal of an attainable region (AR) framework for the analysis of open milling circuits.

Following the successful use of the AR technique, it is fair to state that the method has grown to becoming an alternative tool for the analysis of full-scale milling data.

However, for the tool to be effectively used, one should rely on the classical milling model and build a robust simulator validated against industrial data. Once this is done, simulation data can be generated for analysis and optimization following the methodology proposed in this article.

Future studies will examine the effects of slurry density on milling. Energy usage for ball milling will also be included in the AR analysis. The possibility of generalizing the method to encompass closed milling circuits will also be considered.

Equally important is the holistic integration of downstream concentration processes such as flotation with ball milling.

This, of course, is dependent on a better characterization of the flotation performance of the ore.

Until all the above is addressed, it can be stated that the present exploratory work has demonstrated the suitability of the AR technique for studying open ball milling-circuits.

Acknowledgements The authors are indebted to the University of South Africa (UNISA) and the University of the Witwatersrand for encouraging the collaborative work between the two institutions.

The industrial data used for validation purposes was the product of successful collaboration between Anglo Platinum, Magotteaux, the University of Cape Town, and the University of the Witwatersrand.

Special thanks to the Waterval Mine of Anglo Platinum for granting access to the plant, and also for technical support in setting up the experimental programme.

Appreciation is further extended to Dr Chris Rule, Head of Concentrator Technology at Anglo Platinum, for giving clearance to publish the paper as well as the industrial data collected at the Waterval Mine.

Magotteaux (Pty) Ltd, the developer of the specialized sensor Sensomag® used during the sampling campaign, is also acknowledged for the invaluable data extracted from their sensor. This information was critical in obtaining accurate estimates of grinding ball filling.

References AUSTIN, L.G., JULIANELLI, K., DE SOUZA, A.S., and SCHNEIDER, C.L., 2007.

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CHIMWANI, N., GLASSER, D., HILDEBRANDT, D., METZGER, M.J., and MULENGA, F.K. 2013. Determination of the milling parameters of a platinum group minerals ore to optimize product size distribution for flotation purposes.

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is revisited to include milling parameters such as ball filling and slurry concentration. Once this is done, an improved classification model that allows for milling parameters will be used instead of the particle-size-based classification function in Equation [7]. The direct implication will be a robust simulation model with improved prediction abilities.

This brings us to discuss the effect of the first milling parameter investigated, that is, ball filling.

The observed trend was that increased ball fillings yielded a higher production of m2; however, ball fillings above J = 40% brought about the opposite effect. This is consistent with the observation that wet mills are efficient when slurry completely occupies the interstices between the grinding balls (Latchireddi and Morrell, 2003; Tangsathitkulchai, 2003). In this case, the slurry filling of the mill is equal to unity (i.e. U = 1). Keshav et al. (2011) also reported similar findings, with an improved reduction ratio being noted when ball filling was increased. However, their target was the production of <75 μm and not m2 (i.e. -75 +10 μm) as is the case in the present work.

On another note, Mulenga and Chimwani (2013) were able to demonstrate that for an overflow discharge mill such as the one currently under consideration, a ball filling of J = 35% represents a threshold beyond which the slurry pool disappears. In other words, J = 35% approximately corresponds a slurry filling U = 1. That is ostensibly why in Figure 12 the optimum mill throughput is recorded at ball filling J = 35–40%.

As far as the second milling parameter is concerned, Figure 13 shows that a higher mass fraction of m2 is produced when the solids feed rate is decreased from 400 t/h to approximately 30 t/h. At feed rates less than 30 t/h, the proportion of m2 in the mill product falls sharply. This behaviour could be attributed to the fact that low flow rate implies a longer residence time of particles inside the mill, and therefore higher grinding levels. However, a much lower flow rate translates into a finer grind, which eventually results in particles reporting predominantly to size class m3 (i.e. -10 μm) and not m2, hence, the sudden drop in throughput.

Next, the effect of mill speed on the production of m2 is linked to the change in load behaviour of the mill. Indeed, a low mill speed is synonymous with low ball-to-ball and ballto-ore impact levels inside the mill. Similarly, a very high mill speed incurs more particle centrifuging, and therefore less impact breakage. So, by and large, speeds of the order of φc = 70–85% of critical ensure a high level of impact breakage and consequently a better grind (see Figure 15).

Lastly, based on the simulation results of the effect of ball size (see Figure 16), grinding balls of diameter d = 10 mm produced almost three times the m2 that was recorded with ball diameter d = 50 mm. This is ascribed to the fact that small balls are known to produce a finer grind (Austin et al., 1984; Katubilwa and Moys, 2009; Cho et al., 2013). They will therefore produce a higher fraction of m2 than larger grinding balls. Notwithstanding this, the life span of the small balls is short, and a mixture of balls of different diameters may be a better option. A study is currently in progress aiming at determining the best ball mix for the production of m2.

Application of the attainable region technique to the analysis of a full-scale mill in open circuit CHIMWANI, N., MULENGA, F.K., HILDEBRANDT, D., GLASSER, D., and BWALYA, M.M. 2014a. Scale-up of batch grinding data for simulation of industrial milling of platinum group minerals ore. Minerals Engineering, vol. 63. pp. 100–109.

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CHO, H., KWON, J., KIM, K., and MUN, M., 2013. Optimum choice of the make-up ball sizes for maximum throughput in tumbling ball mills. Powder Technology, vol. 246. pp. 625–634.

GLASSER, D. and HILDEBRANDT, D. 1997. Reactor and process synthesis.

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GUPTA, A. and YAN, D.S. 2006. Mineral Processing Design and Operation: An Introduction. Elsevier, Perth.

HLABANGANA, N., VETTER, D., METZGER, M.J., GLASSER, D., and HILDEBRANDT, D. 2012. Industrial application of the attainable region analysis to a joint milling and leaching process. Proceedings of the VIII International Comminution Symposium – Comminution ’12, Cape Town, South Africa.

Minerals Engineering International, Falmouth, UK. KATUBILWA, F.M. and MOYS, M.H. 2009. Effect of ball size distribution on milling rate. Minerals Engineering, vol. 22, no. 15. pp. 1283–1288.

KATUBILWA, F.M., MOYS, M.H., GLASSER, D., and HILDEBRANDT, D. 2011. An attainable region analysis of the effect of ball size on milling. Powder Technology, vol. 210, no. 1. pp. 36–46.

KELLY, E.G. and SPOTTISWOOD, D.J. 1990. The breakage function; what is it really? Minerals Engineering, vol. 3, no. 5. pp. 405–414.

KESHAV, P., DE HAAS, B., CLERMONT, B., MAINZA, A., and MOYS, M.H. 2011.

Optimisation of the secondary ball mill using an on-line ball and pulp load sensor – the Sensomag. Minerals Engineering, vol. 24. pp. 325–334.

KHUMALO, N. 2007. The application of the attainable region analysis in comminution. PhD thesis, University of the Witwatersrand, Johannesburg.

KHUMALO, N., GLASSER, D., HILDEBRANDT, D., and HAUSBERGER, B. 2008. Improving comminution efficiency using classification: an attainable region approach.

Powder Technology, vol. 187, no. 3. pp. 252–259.

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KHUMALO, N., GLASSER, D., HILDEBRANDT, D., HAUSBERGER, B., and KAUCHALI, S. 2006. The application of the attainable region analysis to comminution.

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KHUMALO, N., GLASSER, D., HILDEBRANDT, D., HAUSBERGER, B., and KAUCHALI, S. 2007. An experimental validation of a specified energy-based approach for comminution. Chemical Engineering Science, vol. 62, no. 10. pp. 2765–2776.

KING, R.P. 2001. Modeling and Simulation of Mineral Processing Systems.

Butterworth-Heinemann, Oxford.

LATCHIREDDI, S. and MORRELL, S. 2003. Slurry flow in mills: grate-only discharge mechanism (Part 1). Minerals Engineering, vol. 16, no. 7. pp. 625–633.

METZGER, M.J. 2011. Numerical and experimental analysis of breakage in a mill using the attainable region approach. PhD thesis, State University of New Jersey, New Brunswick.

METZGER, M. J., DESAI, S. P., GLASSER, D., HILDEBRANDT, D., and GLASSER, B. J. 2012. Using the attainable region analysis to determine the effect of process parameters on breakage in a ball mill. American Institute of Chemical Engineers Journal, vol. 58, no. 9. pp. 2665–2673.

METZGER, M.J., GLASSER, D., HAUSBERGER, B., HILDEBRANDT, D., and GLASSER, B.J. 2009. Use of the attainable region analysis to optimize particle breakage in a ball mill. Chemical Engineering Science, vol. 64, no. 17. pp. 3766–3777.

MULENGA, F.K. and CHIMWANI, N. 2013. Introduction to the use of the attainable region method in determining the optimal residence time of a ball mill.

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RULE, C.M. and ANYIMADU, A.K., 2007. Flotation cell technology and circuit design – An Anglo Platinum perspective. Journal of the Southern African Institute of Mining and Metallurgy, vol. 107. pp. 615–622.

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VERMEULEN, L.A., HOWAT, D.D., and CAMPBELL, Q.P. 1991. The open-circuit production of material finer than 75 μm in open-circuit mills. Journal of the South African Institute of Mining and Metallurgy, vol. 91, no. 10. pp. 363–367. ◆

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Thermogravimetric investigation of macadamia nut shell, coal, and anthracite in different combustion atmospheres

by S.O. Bada*, R.M.S. Falcon*, L.M. Falcon*, and M.J. Makhula†

The combustion and co-combustion behaviour of macadamia nut shell, high-ash coal, and anthracite, along with their blends was studied using thermogravimetry. The reactivities of all samples were analysed in air, oxygen, and CO2 atmospheres and at different heating rates from 10 to 40°C/min. Macadamia shell was found to have a lower ash content, 0.36%, than coal with 27.49% ash. The calorific values were similar, 19.64 MJ/kg and 19.44 MJ/kg respectively. The differential thermogravimetric results indicate that as the heating rates increase the ignition, peak, and burnout temperatures increase significantly, leading to high combustion rates. The interaction between the fuels was evaluated using the weighted average model, and the results indicated that there is more synergetic interaction between 20% macadamia plus 80% coal under oxygen than in air and CO2 atmospheres. The results of the investigation provide the combustion and co-combustion characteristics of various samples and their blends and indicate their combustion compatibilities.

Keywords anthracite, blending, coal, macadamia shell, combustion.

Introduction The global macadamia nut industry is developing at a very rapid rate. South Africa is currently the world’s third largest producer, with a production of about 35 000 t as of 2012 (Department of Agriculture, Forestry and Fisheries, 2013), compared to 840 t produced in 1996. The greatest numbers of the macadamia nut farms in South Africa can be found in Mpumalanga, Limpopo, and KwaZulu-Natal provinces, with Mpumalanga as the largest producing area. There are many factors influencing the increased demand for macadamia nuts, the most important factor perhaps being their nutritional value (California Dried Fruit and Nuts, 2011; MPC, 2014) and their suitability as a feedstock for manufacturing cosmetics, bakery products, nut paste, chocolates, sauces, and ice-cream. As a fuel, macadamia shell has a high heating value ‘as-fired’ of about 20.71 MJ/kg (Vhathvarothai et al., 2013), which could in turn be used in drying the nust, and the shell could also be milled and applied as a fertilizer.

The shell makes up almost 60–70% by weight of the macadamia nut. Considerable amounts of shells are generated, which are considered as a waste by the macadamia The Journal of The Southern African Institute of Mining and Metallurgy

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* School of Chemical and Metallurgical Engineering, Faculty of Engineering and the Built Environment, University of the Witwatersrand, South Africa. † Department of Mineral Processing, Mintek, South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Nov. 2014 and revised paper received Feb. 2015.

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Synopsis

farmers who have to pay a disposal fee to dump the shells in a landfill. Vhathvarothai et al. (2013) showed that the heating value of macadamia shells is higher than most of the coals used in South African power plants. As the demand for electricity increases, there is need to investigate the co-firing potential of this shell, which is abundantly available in Mpumalanga, the major coal-producing region of South Africa. The utilization of biomass as a single source of fuel for combustion is known to face many challenges, such as the low melting point of the ash, which causes fouling and slagging, varied combustion characteristics, and high moisture and volatile matter contents (Phanphanic and Mani, 2011; Bada et al., 2014a). However, the aforementioned problems could be reduced through the coblending of biomass with coal in the correct proportions. Studies have shown that the blending of biomass with coal could lead to a significant reduction in emissions of CO2 (Li et al., 2012), and an increase in the combustion reactivity of coal by decreasing the ignition temperature and reducing burnout time (Moon et al., 2013; Bada et al., 2014b).

Numerous thermochemical conversion technologies such as gasification and combustion have proven to be promising ways of producing energy from both coal and biomass. The use of thermogravimetric analysis (TGA) to investigate the thermal characteristics of macadamia nut shell during combustion may provide an insight into the utilization of this resource for power generation in South Africa. With thousands of boilers in South Africa, the co-firing of macadamia and coal might be an added

Thermogravimetric investigation of macadamia nut shell, coal, and anthracite advantage to those using smaller boilers, by reducing their CO2 emissions and improving their energy output.

Furthermore, with millions of tons of high-ash coal in South Africa, the co-firing of macadamia might be an attractive option to enhance the thermochemical performance of this low-grade coal. In addition, the success of this approach is likely to develop the possibility of using the considerable tonnages of coal fines generated in the country in a cleaner and more efficient manner. As there is little published literature on thermogravimetric investigations of macadamia combustion, this paper seeks to investigate the viability of using macadamia nut shells as a renewable energy source for combustion and co-combustion applications, and to determine the physiochemical and combustion properties of the material and its co-firing potential with high-ash coal.

In this study, the reactivities of macadamia nut shell, high-ash coal, and anthracite, individually and in blended proportions, were analysed thermogravimetrically in air, oxygen, and CO2 atmospheres at different heating rates. The macadamia/coal and macadamia/anthracite blends were evaluated at different ratios using differential thermogravimetric (DTG) techniques.

for all samples using a Leco AC500 calorimeter in accordance with ASTM D5865-04. The system uses an electronic thermometer with an accuracy of 0.0001°C to measure the temperature every six seconds, with the results obtained within 4.5 to 7.5 minutes. Ultimate and sulphur analyses of all samples were performed according to ASTM D 5373-02 and ASTM D 4239-05 for CHN and sulphur content respectively, using a Leco CHN 628 with add-on 628 S module.

The experimental DTG profiles obtained from the cocombustion of the fuels at different weight proportions were used to study the interaction between coal and macadamia nut shell. The theoretical DTG curves of the various blends were calculated by summing the weight loss rates of each individual sample and comparison with the results from the experimental DTG curves. This sought to confirm whether synergistic interactions occurred between the macadamia and coal components in the blends. The theoretical relationship used was defined as follows (Gil et al., 2010; Wang et al., 2011): M = xmaca . Mmaca + xcoal . Mcoal

[1]

Experimental

where xmaca and xcoal are the fractions of macadamia and coal in the blend, respectively, and Mmaca and Mcoal are the weight loss rates (%/min) of the individual fuels in the blend.

Sample preparation

Results and discussion

The macadamia material and high-ash coals were sourced from the KwaZulu-Natal province of South Africa. The shell was milled using a Retsch SM 200 cutting mill to -212 μm.

The anthracite and the coal samples were milled in a hammer mill and then pulverized to -212 μm, with the representative fractions prepared for the combustion and co-combustion investigation. Blends of macadamia/coal and macadamia/anthracite were prepared using 20%, 50%, and 80% macadamia shell by weight. The calorific values and proximate and ultimate analyses of the individual samples were determined and are presented in Table I.

Thermogravimetric analysis The combustion and co-combustion of the macadamia nut shell, coal, anthracite and their blends in different weight ratios, were investigated using a TGA 701 Leco thermogravimetric analyser. The three raw samples were combusted at heating rates of 10, 20, 30 and 40°C/min in an oxygen atmosphere from a temperature of 25°C to 950°C. Thermal combustion and co-combustion of all raw samples and their blends were conducted at a constant heating rate of 10°C/min from 25°C to 950°C under three different atmospheres (air, CO2, and O2) and the samples were held at 950°C until no further weight loss was recorded. For each experiment, 80 mg of fuel was used. The derivative thermogravimetric (DTG) curve generated provided information on the reactivities of the individual samples and blends under these different conditions.

Material characteristics and analysis of sample interaction The proximate analyses for all samples were conducted in accordance with ASTM D-5142, with approximately 1 g of material used in determining the inherent moisture, ash content, and volatile matter present, and fixed carbon calculated by difference. The calorific value was determined

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Characteristics of Coal A, raw macadamia, and thermally treated PA The results of the proximate and ultimate analyses and the calorific values of the macadamia nut shell, coal, and anthracite are depicted in Table I. The ash content of the macadamia shell was found to be 0.36%, compared to the high ash content of coal (27.49%). The ash content of the shell was found to be much lower than that of most woody and non-woody biomass feedstocks (Xiang et al., 2012; Idris et al., 2012; Slopiecka et al., 2012; Bada et al., 2014a, 2015).

The fixed carbon content of the anthracite sample was significantly higher than that of the coal and the macadamia nut shell. The volatile matter and moisture contents of the macadamia were higher than the values for coal. Both samples had similar calorific values. The nitrogen and sulphur contents of the macadamia nut shell were very low at 0.30%, compared with 1.36 % nitrogen and 0.76 %, sulphur in the coal. The lower nitrogen content of the macadamia shell could support its application as a fuel in combustion systems, thereby reducing the emission of NOx in the flue gas when co-fired with coal.

Influence of different heating rates on combustion curves in an O2 atmosphere The DTG curves at different heating rates (10 to 40°C/min) in oxygen for raw macadamia, coal, and anthracite are presented in Figure 1. The characteristic parameters of the fuels, i.e. ignition, peak, and burnout temperature, are shown in Table II. The DTG profile presented two peaks for the raw macadamia at a heating rate of 10 to 40°C/min. The intensity of the shoulders was more pronounced for the macadamia sample at 10°C/min compared to other heating rates (Figure 1a). The shoulder observed in the temperature region of 230–340°C for the curve at 10°C/min could be attributed to the decomposition of the low molecular weight compound The Journal of The Southern African Institute of Mining and Metallurgy

Thermogravimetric investigation of macadamia nut shell, coal, and anthracite Table I

Physico-chemical properties of raw macadamia, coal and anthracite Parameter

Macadamia

Coal

Proximate analysis (wt%, db) Fixed carbon (db) 23.40 20.76 Volatile matter (db) 76.25 51.75 Ash content (db) 0.36 27.49 Moisture (ar) 7.97 6.49 Ultimate analysis (wt%, ar) Hydrogen 6.20 3.53 Nitrogen 0.30 1.36 Carbon 49.70 52.70 Sulphur 0.30 0.71 Oxygen 35.20 9.5 Ash 0.33 25.71 CV (MJ/kg) 19.64 19.44 db: Dry basis; ar: As received; CV: Calorific value; O: Oxygen by difference [100-(H+C+N+Ash+H20+S)]

‘hemicellulose and some cellulose’, with the evolution of gases such as CO and CO2. The minor peak seen in the region of 380–460°C for the curve at 10°C/min can be ascribed to the complete degradation of the organic functional groups of the macadamia nut shell. This observation is in agreement with the literature (Yang et al., 2007; Pickard et al., 2014).

A decrease in the intensity of the second peak was observed as the heating rate increases from 10 to 40°C/min.

This could be a result of the decrease in time for devolatilization to occur and a reduction in reaction residence time. It was also observed that as the heating rate increases, the DTG curves for all samples move into a higher temperature region, with a wider combustion range at higher temperatures.

Anthracite

73.40 5.80 20.80 4.87 1.00 0.85 67.60 0.30 5.60 19.79 23.10

This is considered to be due to the limitation of heat transfer to the inner part of the sample for effective combustion at the higher heating rates. With lower heating rates, more effective heat would be transferred steadily into the inner part of the sample, resulting in more rapid decomposition of the sample (Xie and Ma, 2013). Based on the thermal characteristic data (Table II), it can be seen that as the heating rate increases, the ignition, peak, and burnout temperatures increase significantly in the case of all the samples (macadamia, coal, and anthracite). In addition, an increase in reactivity (%/min) was noted as the heating rate increases, which indicates that combustion intensity is enhanced by higher heating rates (Chen et al., 2012; Xie and Ma, 2013).

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Figure 1 – DTG curves for macadamia nut shell, coal, and anthracite at different heating rates

Thermogravimetric investigation of macadamia nut shell, coal, and anthracite Table II

Thermal characteristics of raw macadamia, coal, and anthracite Samples

Heating rate

IT (FC) °C

PT (°C)

PT wt loss (%)

BT (°C)

(°C/min) Macadamia

10 160.22 320.70 56.30 20 169.00 323.20 59.80 30 171.43 337.74 64.53 40 230.40 346.66 64.93 Coal 10 231.32 390.00 25.10 20 225.10 469.00 51.12 30 264.50 498.22 55.35 40 241.90 571.44 64.61 Anthracite 10 415.00 502.85 27.81 20 428.13 609.52 53.30 30 431.04 645.00 53.50 40 483.41 674.00 52.20 IT: ignition temperature; PT: peak temperature; BT: burnout temperature; DTGmax: reactivity

520.52 556.40 630.43 657.01 632.20 650.00 706.11 753.03 677.80 778.23 847.20 931.40

Final wt loss

DTGmax

(%)

(%/min)

99.90 99.98 99.95 99.52 72.13 72.44 72.60 72.50 80.20 79.85 80.60 80.41

9.11 17.60 21.70 25.00 4.50 6.14 9.03 10.70 6.40 9.75 10.72 10.90

Influence of different atmospheres on combustion at 10°C/min

Co-combustion of different fuel blends at 10°C/min under air, CO2, and O2 atmospheres

The TG and DTG curves for raw macadamia, coal, and anthracite at a heating rate of 10°C/min under three different atmospheres are presented in Figures 2 and 3. All samples followed the same trend of moisture removal. As the temperature increases, ignition of the hemicellulose content of the macadamia occurs at a temperature range of around 160 to 164°C under all three atmospheres. The peaks are seen overlaying each other, with the sample under an oxygen atmosphere igniting first and having the highest reactivity of 9.11%/min (Table II). The second peaks observed for the three samples (Figure 3) show distinctive curves under different atmospheres. This could be attributed to the differences in the chemical composition and thermal characteristics of the organic functional group within the samples and their individual interaction with the gases used. A significant mass loss of 56.30% was obtained for macadamia combusted under an oxygen atmosphere, followed by 44.31%, in CO2 and 47.86% in air . The same trend was observed for coal and anthracite. The combustion of all samples in oxygen occured in a lower temperature region with the highest reactivity and shortest burnout time, which is in line with the findings of Pickard et al. (2014). In regard to the burning profile of anthracite in oxygen, the peak temperature was attained at 502.85°C, while in air and CO2 the peaks were attained at temperatures of 612.52°C and 654.53°C, respectively. Figure 2 also shows the TG profiles for the three samples in 100% CO2 atmosphere. The weight loss profiles for the coal and anthracite in CO2 atmospheres, as well as with their DTG curves (Figure 3), were similar, and different from that of macadamia in a CO2 atmosphere. The thermal degradation of macadamia is completed over a wider temperature range compared to coal and anthracite (Figure 3). In summary, significant differences in fuel mass loss rate, ignition time, and burnout time were seen to be due to the influence of the different combustion atmospheres on each product tested. The differences between the fuel combustibility at ignition and burnout are likely to be a result of compositional differences, such as volatile matter and fixed carbon content, between the samples.

The co-combustion of coal and anthracite with 20, 50, and 80 wt.% macadamia was investigated under three different atmospheres. The thermal profiles obtained from the TGA tests were used to compare the combustibilities of the fuels as shown in Figures 4–6. The thermographs illustrate the high combustibility of the macadamia/coal and macadamia/anthracite blends. Figure 4 shows that the cocombustion of 20% macadamia in the coal blend has the

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Figure 2 – TGA curves for macadamia, coal, and anthracite under air, oxygen, and carbon dioxide

Figure 3 – DTG curves for macadamia, coal, and anthracite under air, oxygen, and carbon dioxide The Journal of The Southern African Institute of Mining and Metallurgy

Thermogravimetric investigation of macadamia nut shell, coal, and anthracite highest reactivity of about 12.50 %/min in oxygen compared to the tests under CO2 and air. The 50% and 80% coal samples in the macadamia blends have a higher reactivity in the lower temperature region than the raw macadamia and raw coal samples. Furthermore, Figure 4 shows that 50% and 20% anthracite blended with raw macadamia combusts in the same temperature range as the raw macadamia. Blending with macadamia shell resulted in a significant improvement in the combustibility of both coal and anthracite, as seen by the decrease in their ignition temperatures. The ignition temperature was decreased from 415°C (raw anthracite) to around 220°C in co-firing with macadamia nut shell.

The DTG profiles presented in Figures 5 and 6 illustrate the thermographs of coal/macadamia and coal/anthracite blends in air and CO2 respectively. In air, raw macadamia nut shell had the highest reactivity of 8.85 %/min and the lowest burnout temperature. This is in contrast to results obtained under oxygen, where 20% macadamia plus 80% coal had the highest reactivity. It was observed that all coal/macadamia blended samples have lower burnout and ignition temperatures in the lower temperature zones compared to coal. In addition, the anthracite/macadamia blends also followed the same trend under the air atmosphere, with 80% macadamia plus 20% anthracite having a higher reactivity and lower burnout temperature than the raw coal. Under the CO2 atmosphere (Figure 6), the sample with the lowest coal ratio

in macadamia/coal blends (80% macadamia plus 20% coal) was seen to have a similar mass loss rate to that of the raw macadamia. In summary, blending with macadamia shells resulted in a significant improvement in the combustion of the high-ash coal and anthracite samples. The second ‘shoulder’ peaks seen for all samples tested decreased as the percentage of coal and anthracite in the blends decreased.

Interaction between the blends of macadamia and coal under air, CO2, and O2 The calculated and experimental DTG curves were compared with the aim of determining the combustion behaviour of macadamia shell and coal when blended at different ratios, and under different atmospheres. The calculated DTG curves were derived from Equation [1] as the weighted average of the individual fuels in the decomposition. The interaction between the blends of macadamia and coal at 10°C/min, i.e. the experimental total weight loss rate (%/min), could be assumed to be additive or non-synergistic if it is equal to the calculated weighted average derived from Equation [1]. The calculated and experimental DTG curves for 20% macadamia and 50% macadamia blends with coal at a heating rate of 10°C/min are depicted in Figure 7. The calculated and experimental DTG curves for 20% macadamia plus 80% coal under all atmospheres show good agreement at the initial and burnout stages. Significant deviation of about 12%/min was noted for the calculated and experimental 20% macadamia plus 80% coal sample under oxygen during the main massloss stage at a temperature above 240°C. The experimental weight loss rate (%/min) lagged behind the calculated weight loss rate, and the DTG curve was shifted to a lower temperature. The deviations observed for samples under air and CO2 atmospheres were not as significant as that under oxygen, indicating there is more synergetic interaction between macadamia and coal under an oxygen atmosphere.

As the percentage of macadamia shell in the blend increases to from 20% to 50%, as shown in Figure 8, the deviation in the experimental and calculated DTG curves of the samples in air and CO2 atmosphere increases (compare with Figure 7). The ignition, peak, and burnout temperatures for the experimental DTG curves for the three samples are

Figure 4 – DTG profiles of coal/macadamia and coal/anthracite blends in oxygen at a heating rate of 10°C/min

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Figure 6 – DTG profiles of coal/macadamia and coal/anthracite blends in CO2 at a heating rate of 10°C/min VOLUME 115

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Figure 5 – DTG profiles of coal/macadamia and coal/anthracite blends in air at a heating rate of 10°C/min

Thermogravimetric investigation of macadamia nut shell, coal, and anthracite shifted to lower temperatures. From a qualitative point of view, the experimental curves exhibit significant deviations from the theoretical curves, particularly in the temperature range of 240–630 °C. This indicates that the interaction between macadamia and coal is more pronounced at higher temperatures. It is suggested that the deviation between the calculated and experimental DTG curves at higher temperature could be a result of the gases and heat released during the decomposition of the macadamia, which accelerated the combustion of coal in the blend.

Conclusion The results suggest that macadamia shells may constitute a valuable and efficient source of biomass for co-firing with coal and anthracite. The thermographs obtained from the cocombustion of macadamia/coal blends at 20%, 50% and 80% weight percentage macadamia in oxygen showed that all the blends have higher reactivities than coal alone, and that the high reactivities occur in the lower temperature regions.

Combustion in air is slightly less efficient, but may be more economical than in oxygen. Anthracite combusts in a much higher temperature range than coal and macadamia, and therefore may constitute a long-term compatible blend component.

Acknowledgements The authors gratefully acknowledge the financial support of the NRF SARChI Clean Coal Technology and the University of the Witwatersrand. We thank the Mineral Processing

Department, Mintek for permission to publish and for access to some of their research facilities.

References BADA, S.O., FALCON, R.M.S., and FALCON, L.M. 2014a. Investigation of combustion and co-combustion characteristics of raw and thermal treated bamboo with thermal gravimetric analysis. Thermochimica Acta, vol. 589. pp. 207–214.

BADA, S.O., FALCON, R.M.S., FALCON, L.M., and BERGMANN, C.P. 2014b. Co-firing potential of raw and thermally-treated Phyllostachys aurea bamboo with coal. Energy Sources, Part A. (In press).

BADA, S.O., FALCON, R.M.S., and FALCON, L.M. 2015. Characterization and cofiring potential of a high ash coal with Bambusa balcooa. Fuel, vol. 151. pp. 130–138.

CALIFORNIA DRIED FRUIT AND NUTS. 2011. Nut Crop Report, 2011. http://www.stapleton-spence.com/crop-reports/june-2011-nut-cropreport/ [Accessed 8 September 2014].

CHEN, C.X., MA, X.Q., and HE, Y. 2012. Co-pyrolysis characteristics of microalgae Chlorella vulgaris and coal through TGA. Bioresource Technology, vol. 117. pp. 264–273.

DEPARTMENT OF AGRICULTURE, FORESTRY AND FISHERIES, SOUTH AFRICA. A profile of the South African macadamia nuts market value chain, 2013. Pretoria.

GIL, M., CASAL, D., PEVIDA, C., PIS, J., and RUBIERA, F. 2010. Thermal behaviour and kinetics of coal/biomass blends during co-combustion. Bioresource Technology, vol. 101, no. 14. pp. 5601–5608.

IDRIS, S.S., RAHMAN, N.A, and ISMAIL, K. 2012. Combustion characteristics of Malaysian oil palm biomass, sub-bituminous coal and their respective blends via thermogravimetric analysis (TGA). Bioresource Technology, vol. 123. pp. 581–591.

LI, J., BRZDEKIEWICZ, A., YANG, W., and BLASIAK, W. 2012. Co-firing based on biomass torrefaction in a pulverized coal boiler with aim of 100% fuel switching. Applied Energy, vol. 99. pp. 344–354.

MACADEMIA PROCESSING COMPANIES (MPC) Ltd. 2014. Macademia kernel in various market segments. http://www.mpcmacs.com.au/wholesaleproducts.html [Accessed 28 August 2014].

MOON, C., SUNG, Y., AHN, S., KIM, T., CHOI, G., and KIM, D. 2014. Effect of blending ratio on combustion performance in blends of biomass and coals of different ranks. Experimental Thermal and Fluid Science, vol. 47. pp. 232–240 PHANPHANICH, M. and MANI, S. 2011. Impact of torrefaction on the grindability and fuel characteristics of forest biomass. Bioresource Technology, vol.

102, no. 2. pp. 1246–1253.

PICKARD, S.C., DAOOD, S.S., POURKASHANIAN, M., and NIMMO, W. 2014. Co-firing coal with biomass in oxygen- and carbon dioxide-enriched atmospheres for CCS applications. Fuel, vol. 137. pp.185–192.

Figure 7 – Calculated and experimental DTG curves for 20% macadamia blends under different atmospheres

SLOPIECKA, K., BARTOCCI, P., and FANTOZZI, F. 2012. Thermogravimetric analysis and kinetic study of polar wood pyrolysis. Applied Energy, vol. 97. pp. 491–497.

VHATHVAROTHAI, N., NESS, J., and YU, J. 2013. An investigation of thermal behaviour of biomass and coal during co-combustion using thermogravimetric analysis (TGA). International Journal of Energy Research, vol. 38, no. 6. pp. 804–812.

WANG, Q., ZHAO, W., LIU, H., JIA, C., and LI, S. 2011. Interactions and kinetic analysis of oil shale semi-coke with cornstalk during co-combustion.

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XIANG, F. and LI, J., 2012. Experimental study on ash fusion characteristics of biomass. Bioresource Technology, vol. 104. pp. 769–774.

XIE, Z. and MA, X. 2013. The thermal behaviour of the co-combustion between paper sludge and rice straw. Bioresource Technology, vol. 146. pp. 611–618.

Figure 8 – Calculated and experimental DTG curves for 50% macadamia blends under different atmospheres

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YANG, H., YAN, R., CHEN, H., LEE, D.H., and ZHENG, C. 2007. Characteristic of hemicellulose, cellulose and lignin pyrolysis. Fuel, vol. 86. pp. 1781–1788. ◆ The Journal of The Southern African Institute of Mining and Metallurgy

http://dx.doi.org/10.17159/2411-9717/2015/v115n8a11

Application of systemic flow-based principles in mining

by J.O. Claassen*

Research approach Mining value chains are dynamic systems and should be managed according to systemic flow-based principles. This includes a detailed focus on the impact of variable geological conditions on material flow, product quality, and production cost, visibility of the total mining system to operators and managers, as well as the synchronization of resources and activities in downstream processes. Thirty mining operations across the African continent were studied to establish to what extent systemic flowbased principles are applied in day-to-day operations. The study indicated that although most mining operations have identified the relevant geoprocessing variables, only a small number apply this information in a flow-based management approach. Most mining operators do not focus on developing a clear flow view of the mining value chain and on making it visible to operators and managers to enable them to optimally set and reset operations. Daily synchronization of geoprocessing variables is limited to a small number of variables, e.g. synchronizing ore hardness with drilling or milling requirements, in less than 30% of the operations evaluated. Operations where these principles have been adopted reported a noteworthy improvement in performance. The study indicates that notable potential exists in most of the operations evaluated to implement flowbased management principles, including geometallurgical principles. This can significantly improve value chain performance without exorbitant capital layouts and enhance ROCE. Keywords geological variables, mining value chain, mining systems, mining management, systems management, systemic, synchronization, geometallurgy.

Introduction Objective of the study The study seeks to establish to what extent systemic flow-based principles have been adopted in the southern African mining industry and the implications thereof. Specific attention is given to the importance of defining and managing geoprocessing dependencies (geological variables directly impacting mining and plant performance – the geometallurgical approach), establishing a flow view of the mining system, which typically includes a dashboard of the flow of ore/product quality and cost from the mining face to the customer, and limiting the impact of dependencies through synchronization of activities and resources, e.g. ensuring that a continuous miner does not wait for a roofbolter, or that reagent addition rates are matched with a specific ore/concentrate composition.

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On-site research was performed during the period 2009–2013 at 30 mining operations on the African continent (mostly from southern Africa), including small- to large-scale operations and different commodities. The research focused on the following areas: ➤ The impact of geology (ore and orebody morphology) on the performance of the mining system as a whole ➤ Development of orebody domaining logic for a mine and an indication of how this logic can be used to synchronize ore and orebody characteristics with downstream processes ➤ Managing geological and resource requirements to comply with a holistic and systemic approach towards mining ➤ Development of performance and throughput logic for selected geological variables at a mining operation ➤ The influence of two non-compatible material types (material types that behave differently during processing, e.g. soft ore and hard host rock) on mining and plant processing performance and product value.

Systemic flow-based principles were employed to study these areas and to compile the maturity matrix shown in Figure 1. The systemic flow-based principles considered included the following: ➤ Systems comprise elements and parts that are dependent on each other and on the environment; dependencies can be sequential or non-sequential ➤ Material, information, and money flow through the system in a specific direction ➤ Changing properties of material, information, and money as these flow through the system

* University of the Free State, South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Oct. 2014 and revised paper received Apr. 2105.

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Synopsis

Application of systemic flow-based principles in mining ➤ Definition of the system by the interrelationship between the elements, parts, material/information; and its environment ➤ In the case of flow in a system, a capacity-constrained resource (resource with the least capacity) exists ➤ In the case of flow in a system, variability exists ➤ High levels of variability render system performance unstable and unpredictable.

The information obtained was evaluated against requirements defined for the optimal management of complex systems (discussed in ensuing sections). A maturity matrix (shown in Figure 1) was used to score the mining operations studied.

Figure 1 aims to summarize the elements required to establish and maintain a stable and predictable mining environment from a systemic flow-based perspective. The emphasis is on the synchronization of activities and resources, which is discussed in more detail in the following paragraphs.

In addition to this evaluation, a number of industry examples that support the arguments presented in the paper are also included.

Background Mining chains as systems The concepts of value chain and supply chain are well established in the mining industry, e.g. in the logistical systems and spares/consumable management environments.

A supply chain typically consists of components that include people, organizations/departments, infrastructure, flows of material, information flow, and flows of intangible services.

These components combine to improve flow from one area, department, or organization to another for the benefit of all participants (Skyttner, 2001). Similarly, systems are viewed as a composition of finite elements or components, which combine to form an integrated unit that supports a purpose.

Skyttner (2001), Weiss (1971), Boulding (1985), and

Figure 1 – Maturity matrix used for the application of systemic flowbased principles in mining (Laurens, 2013)

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Churchman (1979) all indicated that supply chains (in this case mining chains) have similar characteristics to systems.

This conclusion has far-reaching implications for the management of mining systems when these characteristics are contemplated. These characteristics include (Ackoff, 1981; Backlund, 2000; Carbone, 1999): ➤ Systems are influenced by their environment and other systems. Systems are dynamic, with the interaction of sub-systems across boundaries ➤ Each system component has an effect on how the system functions as a whole ➤ The operation of each system component is interdependent. A network of dependencies and interdependencies exists across system boundaries, which can enhance or ruin system performance.

The implications of each of these points on managing complex mining chains are discussed later in more detail.

Benefits of taking a systems approach The literature elaborates on the benefits of viewing and managing supply chains (mining chains in this case) as dynamic systems, i.e. managing flow in the system as a whole. Cooper et al. (1997) and Stonebraker and Afifi (2004) explain how valuable resources are wasted if supply chains are not adequately streamlined and managed. This situation stems from a lack of integration and knowing which flow constraints determine the system’s output. This in turn directly impacts the allocation of capital and the productivity of resources, which are some of the main concerns of mining at present. Mentzer et al. (2001) also explain how supply chain management can lower costs, improve customer satisfaction (internal and external), and create a competitive advantage for the business. Here the authors emphasize the contribution of understanding and managing all dependencies (variables impacting flow), behaviours, and flows (material/products, information, financial resources, demands and forecasts) at different levels in the organization.

Another important issue in mining systems management is the alleviation of the impact of up– and downstream variability on performance. Since all resources, activities, equipment, and processes are dependent on each other, disturbances ripple downstream and upstream through the production chain if not effectively dampened by buffers.

Forrester (1958), Fowler (1999), Towill (1996), and Wikner et al. (1991) reported that this ripple effect is amplified as it moves away from the source up or down the supply chain, with a significant impact on the performance of the system as a whole due to its destabilizing effect, i.e. increased variability hampers synchronization in the system.

Understanding which factors cause these ripple effects from a flow perspective can limit the impact on supply chains as it assists to optimally set and reset equipment and processes as well as to optimally allocate resources to where they are needed most in the production system. Changes in the geological environment and mining and plant conditions impact not only production, but often also product quality.

Variability in product quality in turn has a serious impact on the performance of customer processes, and every effort should be made to produce a consistent product quality (Everett, 2001).

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Application of systemic flow-based principles in mining

Management of mining systems With the abovementioned system characteristics in mind, some elements of a mining management approach based on flow principles should be highlighted.

Management of the geological environment Most mining operators are acutely aware of the negative impact that changes in the internal and external environment can have on their performance. In the medium to longer term this is typically dealt with by becoming leaner (limiting waste in all parts of the business), employing risk/weight factors to adjust plans and forecasts to make them more ‘realistic’, and developing internal/external relationships that can benefit the organization. On the other side of the spectrum, management of a changing/variable mining environment on a day-to-day basis also poses many challenges. For example, variability in the ore morphology (mineralogy, texture, weathering effects, etc.) and orebody morphology (seam thickness, roof and floor conditions, dip, etc.) from one mining block/area to the next can have a significant ripple/destabilizing effect on downstream processes, as alluded to earlier.

When the management of variability in specifically the geological environment is contemplated from a systemic flowbased perspective, the following requirements should be highlighted: ➤ All the ore and orebody characteristics that impact the flow and quality of the ore, as well as the cost of production per mining area (per block for complex orebodies and per pit/shaft for less complex ores), must be determined. Capacities, efficiencies, recoveries, and costs are then linked to the most important ore and orebody characteristics per area, which are then used in models, mine plans, and forecasts. This is typically the so-called geometallurgical approach (currently focusing mainly on the impact of variable ore characteristics on plant performance and not that much on the impact of variable orebody properties). For example, if variability in ore hardness is a key driver of production performance (production rate, recovery, and cost), then the orebody is classified according to ore hardness categories and mined in a manner that does not compromise the performance of the system as a whole.

With this approach there is a definite move away from considering only average ore grades (Schouwstra, The Journal of The Southern African Institute of Mining and Metallurgy

2010), ore volumes, and averaged correction factors to a more condition-based approach, i.e. the specific environment dictates production performance ➤ Designing and setting up a system at a strategic and tactical level does not suffice to optimize supply chain performance, as unexpected events can occur during operations. In order to counter the impact of these events, an ability to identify and exploit the system constraint(s) and key dependencies in the system on a day-to-day basis is required, i.e. managing system dynamics (ever-changing geological and mining environments, HR dynamics, etc.) on an ongoing basis.

Exploitation of the resource(s) that determines the performance of the system as a whole requires timeous identification and management of all factors, in this case variable geological factors, that impact system performance. Mouritus and Evers (1995) indicated that the most important component in a flow chain is the human component. It can therefore be argued that operators and their managers must have the skills to identify the dynamics in a system (Skyttner, 1996) and set/reset the system to give optimal performance in terms of production rate, ore/product quality, and cost.

This requires an ability to develop a systemic flow view (compared to a functional and process view) of the mining environment, manage the system as a whole, and synchronize activities/cycles, among other things, as illustrated in Figure 1 and discussed in more detail in the following paragraphs.

Management of the system as a whole Figure 2 illustrates that mining personnel can adopt different views of how a system can be viewed and managed. Figures 2a and 2b depict a functional and an activity-based view, respectively. Here the emphasis is on optimizing each part of the system, i.e. a fragmented approach. Figure 2c illustrates the process dependencies in a continuous miner – roofbolter – shuttle car system, each with its inherent performance variability. Operation of this system within an ever-changing geological, mining, and business environment is also shown.

Figure 2 demonstrates that management of a system as a whole not only implies managing the entire system comprising the different functional departments or processes.

It actually implies managing all key dependencies and interdependencies among resources, as well as resources and the environment in order to maintain synchronization in the system (Ackoff, 1981; Backlund, 2000; Carbone, 1999). This requires a detailed systemic flow-based understanding of the operation (refer to Figure 1c), visibility of the entire production chain (in terms of ore/product flow, quality, and costs through the different steps), visibility of the constraint(s), and visibility of triggers/factors (e.g. geological factors) that impact the performance of the constraint(s).

This enables operators, managers, and individuals responsible for the management of the end-to-end process to optimally operate a complex mining system (Mouritus and Evers, 1995).

It should also be noted that mapping out processes (Figure 1b) and creating an understanding of the key processes in an organization should be followed by putting measures in place to manage flow in these processes.

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Furthermore, the number of dependencies and interdependencies (a complex network of dependencies and interdependencies exists) in supply chains is increasing with changes in the internal and external environment and organizational growth. Funk (1995), Mughal and Osborne (1995), Osborne (1993), and Simatupang and Sridharan (2002) all report that increased supply chain complexity is associated with poor performance. Approaching/managing supply chains and systems from a flow perspective can significantly simplify these complex networks and improve business performance.

Finally, managing a supply chain from a flow perspective develops an understanding of the key business drivers of an organization at all levels. This aligns performance expectations by executives and managers with actual supply chain performance. Underperformance issues can then be effectively dealt with at all levels (Skyttner, 2001).

Application of systemic flow-based principles in mining

all practical purposes left to chance, which in turn can create more dependencies and thus increase the complexity in mining systems.

rate and the amount of a competing species in the feed to a flotation circuit ➤ The link between variable orebody morphology, e.g. undulating roof and floor conditions, and equipment set-points has a detrimental impact on mining rates and the amount of dilution in the plant feed. Steepdipping ore zones, faulted areas, changing seam thicknesses, and the presence of dykes have a similar impact on the performance of the mining value chain ➤ The link between different mineralogical entities (e.g. ore vs ore and ore vs waste) and its behaviour through the downstream processes, e.g. a mixture of hard host rock and soft ore or a mixture of fine- and coarsegrained ore. These factors have different influences on downstream processes – the concept is referred to as the compatibility of material. If entities have poor compatibility, the process performance as a whole will be adversely impacted ➤ The dependencies created through compounding effects (different variables impacting a resource simultaneously) are seldom considered. A case in point is when the feed to a dense medium cyclone contains high levels of ultra-fines, a high percentage of neardense material (NDM), and a lower ore to waste ratio than planned. The ultra-fines can typically impact the rheology of the medium, as this changes the apparent viscosity of the dense medium. NDM in the ore results in the so-called cut-point shift, where separation takes place at a higher density than the medium density; and when one of the outlets also becomes a physical flow constraint due to higher waste levels in the feed, the output of the system becomes uncontrollable and unpredictable.

Dependencies are typically managed using time, space, volume, or capacity buffers. A well-placed and managed buffer (Goldratt and Cox, 1998) can limit the impact of dependencies and simplify complex mining systems. The impact of geoprocessing dependencies on the performance of the system as a whole, however, can be reduced only if specific ‘flow properties’ of the ore are synchronized with downstream processing activities and set-points, e.g. the mining of compatible material types from different areas should be synchronized and plant settings must match the plant feed flow and quality properties to ensure optimal throughput, product recovery, quality, and cost.

Management of (inter)dependency

Results and discussion

Figure 2c illustrates a simplified view of some (inter)dependencies between underground coal mining equipment, between the equipment and its environment, and the mining system and the macro environment. In addition to the relatively obvious dependencies illustrated here, a number of less obvious dependencies related to the geological environment also exist. These can include the following geoprocessing dependencies: ➤ The link between ore/waste morphology and equipment/plant settings, e.g. mineralogical composition of the feed influences the hardness of the feed, which directly impacts the crusher settings, mill loads, and feed rates in the comminution circuit.

Another example is the link between reagent dosage

In the previous paragraphs the main characteristics of a system were highlighted, namely that systems are sensitive to changes in their environment, system components impact the functioning of the system as a whole, and system components are interdependent. The implications of this for the management of dynamic mining systems were also briefly highlighted, i.e. the necessity to: ➤ Identify and manage the impact of variable ore and orebody characteristics (geological environment) on production rates, product quality, and cost ➤ Manage the mining system as a whole by introducing visible flow and decision-making triggers in the system ➤ Synchronize and maintain synchronization in the system, which can include the synchronization of

Figure 2 – Views of primary mining value chains: (a) functional view, (b) process view, and (c) systemic flow view (continuous miner–shuttle car–roofbolter system in its respective and combined environments)

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Application of systemic flow-based principles in mining Table I

Identification of geoprocessing variables and their use in a flow management approach Number of operations

Identification of geo-processing variables

Geo-processing information used in a daily flow management approach*

Maturity ranking (level)

Ore/waste morphology Orebody morphology Some variables identified in All variables identified in most Limited in most operations some of the operations of the operations Coal 6 Some variables identified in All variables identified in Limited in most operations most of the operations most of the operations Zn/Pb 2 Some variables identified in Most variables identified in Limited in all operations both the operations both the operations Diamonds 4 Some variables identified in Some variables identified in Limited in all operations some of the operations some of the operations Gold 5 Some variables identified in All variables identified in Limited in all operations most of the operations all the operations Platinum 3 Some variables identified in All variables identified in Limited in all operations some of the operations all the operations Chromium 2 Some variables identified in Some variables identified in Limited in both operations both the operations both the operations Manganese 3 Some variables identified in Most variables identified in Limited in some operations some of the operations all the operations Vanadium 2 Some variables identified in Most variables identified in Limited in both operation both the operations both the operations Mineral sands 1 Some variables identified Some variables identified Limited * Measures and targets exist for geoprocessing variables and are managed in-time based on area-specific conditions (condition-based standards) Iron ore

4

activities, ore flow properties with equipment settings, etc., in order to limit the potential negative impacts of (inter)dependencies in the system and across system boundaries.

Different mining operations were evaluated to specifically establish whether geological variables that impact flow in the mining system have been identified and used in a holistic end-to-end management approach that supports synchronization between the characteristics of the ore and equipment settings. The results obtained are discussed in the following section.

Influence of the geological environment Table I indicates (using the maturity matrix depicted in Figure 1) to what extent ore and orebody characteristics/morphology have been identified and used in a flow management approach in the different mining operations evaluated.

From Table I it is evident that most mining operators have identified at least some geoprocessing variables, i.e. geological variables that directly impact mining and plant performance. In fact, some mining houses are actively focusing on geometallurgy to enhance their knowledge of the processing potential of their deposits.

The use of geoprocessing information for the daily management of operations is, however, limited to mining plans â&#x20AC;&#x201C; in some cases it is not used at all. In these operations, the focus is mainly on maintaining average grades, volumes, qualities, and process efficiencies, i.e. system dynamics caused by the presence of constraints and dependencies in the system and across system boundaries receive little attention.

This approach is in most cases supported by a fragmented or functional approach to all aspects of mining (Figure 1a).

However, when an integrated approach is employed in mining and an emphasis is placed on the correctly identified geoprocessing variables, significant financial benefits can be obtained as discussed in the case study below.

The Journal of The Southern African Institute of Mining and Metallurgy

2-3 3-4 3 2 3 3 2-4 2-4 3 3

Case study: management of geoprocessing variables at a mid-sized coal mine Claassen (2013) stressed the importance of identifying and managing geoprocessing variables at a mid-sized coal mine operating in the Witbank coalfields, South Africa. It was shown that blending non-compatible coal and the treatment of ROM material containing high and variable levels of ultrafines, near-dense material, and dilution had a detrimental impact on the performance of the dense media separation (DMS) plant operations. The ability to link geological properties of the ore and orebody such as weathering, undulating floor conditions, and ore texture with downstream processing requirements resulted in the development of condition-based planning standards and improved synchronization of the total system. This in turn improved system stability and rendered operations more predictable. Figure 3 illustrates an almost 30% improvement in DMS plant production performance subsequent to the implementation of flow-based principles at the mine.

Figure 3 â&#x20AC;&#x201C; Improvement in DMS plant performance subsequent to the implementation of flow-based principles at an Mpumalanga coal mine VOLUME 115

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Commodity

Application of systemic flow-based principles in mining Management of the system as a whole The visibility of the primary mining value chain, including key secondary chains (e.g. spares and consumable supply chains), at the different mines selected was evaluated. More specifically, the study aimed to establish to what extent a detailed flow view is created by the mines to assist their operators and managers to manage flow in the system. Table II summarizes the findings of this part of the study (using the maturity matrix depicted in Figure 1).

Some mining operators have developed a detailed flow view (refer to Table II) of parts of their mining systems to assist mainly with the management of logistics, i.e. ore flow to plant or product to market flows. Apart from this, very few mining operators attempt to create a flow view of the end-to-end mining system to assist operators and managers to better operate and manage these systems. One of the chrome mines reviewed, however, developed a simplified flow view of mining operations that yielded a notable improvement in overall performance as discussed in the following paragraph.

Case study: establishing visibility of ore flow at an underground chrome mine A flow view was established and constraints determined for an underground chrome mine consisting of five sections, as illustrated in Figure 4. Prior to a study and implementation of subsequent corrective actions, the system was unstable with constraints frequently moving from one section to another and/or one resource to the next in a specific section. Making the overall mining system visible to operators and managers assisted them to identify the flow constraints, put corrective actions in place, and actively manage the system as a whole.

The intervention implemented resulted in a sustainable increase in production from about 3500 t/d to 4500 t/d (Kloppers, 2013).

Synchronization in mining systems Synchronization of activities, cycles, and the characteristics of the ore with equipment and process settings can be employed to limit the impact of (inter)dependencies in complex supply

chain systems such as mining systems. It could be argued that synchronization should be a key component of a flowbased management approach towards these systems. A focus on synchronization should also be an indicator of whether a mining operator understands mining system flow dynamics and has the ability to apply flow-based management principles. Table III indicates whether the mining operations studied focus on the use of synchronization to limit the impact that (inter)dependencies, specifically geoprocessing dependencies, can have on operational performance.

Table III suggests that only a small number of mining operations, namely 8 out of 30 mines (27%) in this case, focus partially on synchronization in daily operations to limit the impact of (inter)dependencies between geoprocessing variables. This could imply that only a small percentage of mining operators have developed a flow-based understanding of mining systems and are therefore able to optimally exploit their resources, which include the mineral resource(s), equipment, people, and capital (Cooper et al., 1997; Stonebraker and Afifi, 2004).

If it is argued that the mining value chain can be viewed as a dynamic system and its performance is highly dependent on system stability and predictability, it follows that a focus on synchronization of geoprocessing variables can result in significant improvements in business performance, as discussed in the following paragraph.

Case study: synchronizing ore flow characteristics with smelter requirements at a steel producer A steel production value chain comprising coal and iron mines, the respective washing plants, and the smelter was evaluated and solutions implemented to improve synchronization in the system (Laurens, 2007). The aim of the exercise was to improve overall (total value chain) business performance with an emphasis on improvements in net profit and return on investment.

Synchronization between smelter feed material and the operational requirements of the smelter was achieved through the production of coal and iron ore at a consistent quality and rate in the respective washing plants. Consistency in raw material quality and volume at the washing plants, which

Table II

Total mining value chain visibility to operators and managers Commodity

Number of operations

Iron ore

4

Coal

6

Zn/Pb

2

Diamonds Gold Platinum Chromium

4 5 3 2

Manganese

3

Vanadium Mineral sands

2 1

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• Detailed flow view created for parts of the value chain in most operations • Visibility fragmented mostly along functional boundaries in most cases • Detailed flow view created for some mining equipment in some operations • Visibility fragmented mostly along functional boundaries in all cases • Detailed flow view not created • Visibility fragmented mostly along functional boundaries in both cases • Detailed flow view not created • Detailed flow view not created • Detailed flow view not created • Detailed flow view established in one of the operations • Visibility fragmented at one of the operations • Detailed flow view established at one of the operations • Visibility fragmented mostly along functional boundaries at two of the operations • Detailed flow view not created • Detailed flow view created for parts of the operation • Visibility fragmented along functional boundaries

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Maturity ranking (level)

3-4 3-4 2-3 3 3 3 2-4 2-4 2 4

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Application of systemic flow-based principles in mining

Figure 4 – Simplified flow view of five underground sections at a chrome mine (Kloppers, 2013)

Table III

The use of synchronization to minimize the impact of (inter)dependencies between geological variables and processing capabilities in daily operations Commodity

Number of operations

Geo-processing dependencies synchronized in daily operations*

Maturity ranking (level)

Iron ore

4

Limited to some dependencies in one of the operations

2-3

Coal

6

Limited to some dependencies in some of the operations

2-4

Zn/Pb

2

Limited to some dependencies in one of the operations

2-3

Diamonds

4

None

2

Gold

5

None

2

Platinum

3

None

Chromium

2

Limited to some dependencies in one of the operations

2-4

Manganese

3

Limited to some dependencies in one of the operations

2-4

Vanadium

2

None

2

Mineral sands

1

Limited to some dependencies

3

2

* Does not include synchronization in logistical activities such as ore/product transport

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stability followed, which resulted in financials gains ‘beyond expectations’.

Conclusions Research into the application of systemic flow-based principles at 30 mining operations on the African continent indicated that: VOLUME 115

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were managed as the constraints in the system, was in turn achieved through optimal synchronization of plant settings and ore (coal and iron ore) characteristics. The latter was achieved through classification (domaining) of the respective orebodies based on physical and chemical ore flow characteristics and batch-washing the different domains through the respective plants. A significant improvement in system

Application of systemic flow-based principles in mining ➤ Even though geoprocessing variables are identified in a number of operations, they are not actively used in a daily flow management approach at most of the operations ➤ The development of a detailed flow view of operations is mostly limited to areas where logistics forms a key component of mining systems ➤ Visibility of the mining system is mostly fragmented along functional/departmental boundaries; highly fragmented operations do not support flow in the system ➤ Daily synchronization, a key indicator of the application of systemic flow-based principles, is employed at a small number of operations to limit the impact of (inter)dependencies in variable mining systems ➤ A focus on geoprocessing variables, visibility of mining value chains, and the synchronization of mining activities can yield notable improvements in performance as indicated through a number of case studies A thorough understanding of mining system dynamics and an ability to implement systemic flow-based principles at all organizational levels should enable operators and managers to optimally set and reset operations to achieve optimal flow of material, ore/product quality, and costs.

Furthermore, prioritization and allocation of resources (equipment, people, and capital) should be based on flow requirements in the systems. A focus on these factors will enhance synchronization and therefore product production at the required quality and lowest possible cost.

The study indicates that significant potential exist in some of the operations evaluated to implement flow-based management principles, including geometallurgical principles.

This can significantly improve value chain performance without exorbitant capital layouts and enhance ROCE.

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and Metallurgy, vol 113, pp. 761–768.

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dynamics. International Journal Production Economics, vol. 22, no. 3. pp. 231–248.

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http://dx.doi.org/10.17159/2411-9717/2015/v115n8a12

Evaluating the coal bump potential for gateroad design in multiple-seam longwall mining: a case study

by X. Wang*, J. Bai*, W. Li*, B. Chen*, and V.D. Dao*†

This study proposes a methodology for evaluating the risk of coal bumps in multiple-seam longwall mining. Both the stress field and the total energy release (TER) during retreat were evaluated in the specified case involving multiple-seam mining using the LaModel program. The results of numerical simulations indicated that both the peak vertical stresses on the panel edges and the peak TERs in the outby longwalls increased significantly as the horizontal offsets were reduced from 60 m to zero. With the comprehensive consideration of the stress field and TERs, a conservative offset of 60 m was ultimately adopted when developing the gateroads of the lower panel in the field. The field measurements indicated that coal bumps were avoided completely by employing the proposed design, and the maximum roof-to-floor and rib-to-rib convergences of the tailgate during retreat were only 360 mm and 576 mm, respectively.

Keywords coal bump; total energy release (TER); longwall mining; multiple-seam

Introduction Coal bumps – sudden, violent bursts of coal – pose a serious threat to the safe excavation of coal. Their occurrence depends on the properties of the surrounding rocks, stress field, and dynamic disturbance. The complexity of the mechanism of coal bumps is further enhanced in multiple-seam mining due to seam interaction. Multiple-seam mining is most frequently found in the coalfields in eastern China, such as Xuzhou, Xinwen, Zaozhuang, and Huainan. In the past, due to issues of availability, economics, and ignorance of the potential risk in multipleseam mining, coal seams in such conditions were mined without proper planning to account for seam interaction. Fortunately, studies published over the past three decades have raised awareness of the problems associated with multiple-seam mining, and indicated that the ground control issues induced by multiple-seam mining can be avoided and/or minimized with proper planning.

A number of technical studies pertaining to the design of multiple-seam mines have been published. These studies can be divided into two groups based on their methodologies of design: empirical (Chekan and Listak, 1993; Haycocks and Zhou, 1990, Matetic et al.; 1987a, 1987b) and numerical approaches (Li The Journal of The Southern African Institute of Mining and Metallurgy

Geological background Zhuang Shuanglou coal mine is located in Xuzhou, Eastern China. Panels 7111, 7113, and 7115, which are flat and 3.8 m thick, were extracted from the No. 7 coal seam, whereas panels 9111 and 9113, also flat and 3.7 m thick, were extracted from the No. 9 coal seam.

* School of Mines, China University of Min &Tech, China. † Hanoi University of Mining and Geology, Vietnam. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Aug. 2014 and revised paper received Jun. 2015.

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Synopsis

et al., 2000; Maleki, 2007; Mark, 2007; Peng, 2007; Zipf, 2005). The majority of the extant studies on the interaction mechanisms in multiple-seam mining have focused on the load transfer during operation, resulting in the proposal of the pressure bulb theory (Forrest et al., 1987; Haycocks and Karmis, 1983; Peng, 2008; Peng and Chandra, 1980) and pressure arch theory (Holland, 1973; Oram and Ponder, 1997) to guide multiple-seam mine design.

However, neither, of these theories can be used to evaluate the potential for coal bumps.

Recent studies have indicated that coal bumps might occur when inappropriate designs are employed in multiple-seam mining (Peng, 2008). Past studies suggest that a great amount of energy is more likely to accumulate in pillars in the upper coal seam, and disturbances from drives and stopes in lower seams enhance the incidence of sudden breaks, which may also result in a coal bump. Thus, research focused on load transfer in multiple-seam mining cannot effectively prevent the potential bump from occurring. Instead, reasonable operation of roadways is an effective approach to weakening the accumulation of energy in pillars and the disturbance from drives and stopes in lower seams, thus reducing the risk of coal bumps in multiple-seam mining.

This paper evaluates the risk of a coal bump in multiple-seam longwall mining based on numerical modelling of the total energy release (TER).

Evaluating the coal bump potential for gateroad design in multiple-seam longwall mining Table I

Stratigraphic column for the No. 7 and No. 9 coal seams in the Zhang Shuanglou Mine Rock type

Shale No.7 coal seam Shale Fine sandstone No.9 coal seam Shale

Thickness, m

Overburden depth, m

Uniaxial compressive strength, MPa

Young's modulus, GPa

Poisson's ratio

2.13 3.8 0.89 14.11 3.7 1.8

769.87 772 775.8 776.69 790.8 794.5

28.3 6.2 24.4 68.43 6.2 23.7

8.46 2.07 8.32 18.45 2.07 7.86

0.28 0.31 0.29 0.23 0.31 0.29

Geological surveys showed that both the No. 7 and No. 9 coal seams were uniform and their average overburden depths were 772 m and 787 m, respectively. According to the core logs from exploration in these areas, shown in Table I, the interburden of the No. 7 and No. 9 seams was 15 m thick and consisted of a 0.89 m thick shale underlain by a 14.11 m thick fine sandstone. The mechanical properties of these layers, shown in Table I, refer to laboratory test results.

All three panels in the No. 7 seam were mined out and sealed by the end of 2010; the panel layouts are shown in Figure 1. In the beginning of 2012, the same panel layout as 7111 was proposed for panel 9111 (see Figures 1 and 2) to begin the development and retreat in No. 9 seam; however, a slight bump occurred at the pillar 20 m outby the longwall face during panel retreat. Strong pillar spalling and floor heave were observed, and the averaged roof-to-floor convergence reached 1.2 m. Due to safety concerns, panel 09111 was subsequently abandoned. Fortunately, this accident did not result in any fatalities.

Another mine, the San Hejian coal mine, was extracting the No. 7 and No. 9 seams to the southeast of Zhuang Shuanglou. The two mines are separated by a normal fault, with the San Hejian mine on the footwall side. The average fault offset in a vertical direction is 128.5 m. Hence, the overburden depths of the flat No. 7 and No. 9 coal seams in San Hejian Mine were 643.5 m and 658.5 m, respectively.

The panels of the No. 7 and No. 9 seams in this mine were in

the same layout and the upper longwalls were superimposed precisely on the lower ones. Note that both seams in the blue highlighted area in Figure 1 were mined out by 2007. No bump occurred in this mine during the extraction process.

The differences in the ground response between these two mines during operation may thus be attributed to the different overburden depths.

Knowing the lower seam was mineable in the San Hejian mine, the owners of the Zhang Shuanglou mine proposed attempting to extract the No. 9 seam by arranging appropriate horizontal offset of the longwall panels. A schematic diagram of the idea to offset the lower panels is shown in Figure 2.

The problem thus became one of finding the proper offset to avoid potential coal bumps.

Numerical modelling LaModel program The boundary element model (BEM) program LaModel was employed to evaluate the stress distribution and TERs when the lower seam was extracted. LaModel was originally written in 1994 by Dr Keith Heasley (1998), and experience showed that it could be used for structural mine planning and analysis (Heasley et al., 2003; Stemwart et al., 2006). In 2009, energy release rate calculations were incorporated into LaModel 3.0, and a number of case studies indicated that the energy release rate calculation module was fairly reliable for evaluating mine design in bump-prone areas (Sears and Heasley, 2009). Note that LaModel assumes a frictionless coal/rock interface, which may be an unwarranted assumption for coal bump prediction. However, in the present study the model was carefully calibrated to offset this disadvantage.

Model details For the LaModel simulation, the coal seam was discretized with 2 m2 elements in a 250 × 500 element grid with the model boundary shown in Figure 1. The thickness and Young’s modulus of overburden rock in LaModel need to be calibrated before reasonably accurate results can be produced.

Figure 2 – Cross-section A–A (Figure 1) of studied area

Figure 1 – Panel layouts of studied area

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Evaluating the coal bump potential for gateroad design in multiple-seam longwall mining

[1] where Ws is the width of the side abutment (or influenced zone) in feet, and h is the overburden depth in feet.

Note that the US unit used in above equation was mathematically converted into SI. Since the overburden depth of Panel 7111 was 772 m, the width of the side abutment after panel retreating would be 143 m. The rock mass in the overburden was simulated with a modulus of 20.7 GPa and laminations 10 m in thickness. The profile of the side abutment pressure predicted by LaModel is shown in Figure 3. The data indicates that zone of influence of the side abutment pressure reaches 148–150 m, which agrees well with the result calculated using Equation [1]. From this point of view, the selected numbers for the Young’s modulus and thickness of rock mass in the overburden were acceptable.

On the other hand, the strain-softening material model was used for both coal seams, and the element strengths were determined using the in-situ coal strength of 6.2 MPa in conjunction with the Mark-Bieniawski pillar strength formula as implemented in the coal material ‘wizard’ in LaModel (Heasley and Agioustantis, 2007). The coal elastic modulus was set at 2.07 GPa, and the residual seam stress and strain values of coal materials were approximated based on the method of Karabin (1994). In addition, the gob was assumed to be a strain-hardening material and its stress-strain relationship followed Salamon’s gob model equation (1990).

Rigid seam boundary conditions were implemented on the north, south, and west sides of the grid, while a symmetrical boundary condition was implemented on the east side.

The remaining barrier pillars between the adjacent two panels in the upper seam were 15 m wide. The interburden of the two seams was 15 m thick, and all of the longwall panels were 180 m wide, which was determined by mine regulations. A total of eight plans were proposed for the horizontal offsets of panel 9113 (see Figure 2 and Table II).

Figure 3 – Profile of the side abutment pressure of panel 7111 predicted

by LaModel The Journal of The Southern African Institute of Mining and Metallurgy

Table II

Several plans conducted in numerical modeling Plan no.

1

2

3

Offset distance*, m 60 50 40 * Offset distance was shown in Figure 2

4 30

20

6

7

8

10

5

0

The development and retreat of the upper three longwalls were run as the first to sixth steps in the LaModel simulation, followed by development and retreat of panel 9113 with strategies no. 1 through 8 in turn, as shown in Figure 4.

Model results Vertical stress distribution The eight plans were performed separately in turn in the LaModel simulation. Figure 5 shows the vertical stress distribution in 3D during the retreat of panel 9113. Clearly, in Figure 5(a), the offset is 60 m, i.e., the headgate and tailgate of panel 9113 are developed beneath the upper gobs, which may lead to less vertical stress (54.1 MPa,) distributed on the panel edges. In addition, because the interburden is only 15 m thick and the upper remaining pillars may cave down as the longwall 9113 retreats, there is no stress concentration in the middle of the panel 9113 gob.

If the offset distance is equal to zero, however, as shown in Figure 5(b), the maximum vertical stress distributed at the edges of the gob may reach 70.9 MPa. Meanwhile, because the headgate and tailgate are located beneath the upper remaining pillars, the maximum vertical stress distributed at the edges of the entries in the outby longwall will reach 39.4 MPa. This is not a desirable outcome given that the in-situ vertical stress is only 19.6 MPa.

Empirical criterion for TER When operating panel 9113 with different offsets, the TER must also be evaluated because a minor coal bump occurred in panel 9111. It is accepted that in order to prevent bumps from occurring, the TER should be smaller than the minimum energy release that may lead to a coal bump. However, there is no consensus among researchers (Dou and He, 2001), on the ideal value of the limited energy release for specific mining conditions. Fortunately, the San Hejian mine has safely and successfully mined both seams with lower overburden depths due to the fault throw. Thus, the empirical methodology can be used in this case to propose the TER criteria for coal bump control.

Figure 4 – Mining steps of the LaModel simulation

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Due to the difficulty of measuring the distribution of longwall abutment pressures, the following empirical equation (Peng, 2008) illustrating the extent of the side abutment (or the influenced zone) after retreating of panel 7111 was used to calibrate to the inputs in LaModel.

Evaluating the coal bump potential for gateroad design in multiple-seam longwall mining Appropriate offset The peak TERs and predetermined criteria for bump control in panel 9113 in the Zhang Shuanglou mine are plotted in Figure 8. Evidently, peak TERs decrease as the offsets of panel 9113 increase; more importantly, the offsets of panel 9113 need to be at least 50 m to prevent coal bumps according the empirical criteria proposed in the previous section.

Hence, to be more conservative, an offset of 60 m was finally adopted when developing the gateroads of panel 9113 in the field.

Field measurements The roof-to-floor and rib-to-rib convergences in the tailgate of panel 9113 were monitored during the retreat, as shown in Figure 9. It can be seen that (1) the maximum roof-to-floor and rib-to-rib convergences were only 360 mm and 576 mm, respectively; (2) 90% of the entry convergences occurred within the 40 m outby distance to the the longwall face; (3) no violent entry convergence occurred during the longwall retreating. Hence, the proposed design for the gateroad layout was appropriate and successfully avoided a coal bump.

Conclusions

Figure 5 – Vertical stress distribution when operating panel 9113 with different offsets

The numerical model employing LaModel 3.0 was developed as the longwalls retreated successfully in the San Hejian mine. Figure 6 shows the TER after the longwall retreat. Note that the peak TER is 34.8 MJ and is located at the centre of the panel. Experience shows that there will be no bumps under these conditions, therefore in this study 34.8 MJ is selected as the conservative criterion for the offset of longwalls needed to keep the TER less than 34.8 MJ after retreat.

The majority of prior studies of the interaction mechanisms in multiple-seam mining focused mainly on load transfer during operation. However, recent studies have indicated that coal bumps may occur when the longwall layouts in multipleseam mining are improperly designed. Hence, in this study, the boundary element model (BEM) program LaModel was employed to evaluate the stress analysis and total energy release (TER) in the specified case when the longwalls of the lower seam were extracted.

A total of eight strategies were proposed for the horizontal offsets of panel 9113. The model results indicated that the peak vertical stress on the edge of panel 9113 might drop from 70.9 MPa (zero offset) to 54.1 MPa (offset 60 m).

In addition, the TER criterion for bump control in this study was set at 38.4 MJ based on the empirical approach. The TER analysis indicated that the peak TERs decreased as the offsets of panel 9113 increased; more importantly, the offsets of panel 9113 needed to be at least 50 m to prevent coal

TER of different offsets Figure 7 shows the TER when retreating panel 9113 with different offsets. It can be seen that: ➤ There are two peak TERs distributed at the two sides of each entry, and the TERs gradually increase as the offsets decrease from 60 m to zero ➤ When the offsets are less than 40 m, the TERs at the panel edges are greater than those at the panel centre, indicating that the upper remaining pillars will significantly influence the gateroads as the offsets decrease ➤ There is always a peak TER in the centre of panels, no matter what the offset ➤ The peak TERs are 24.3, 28.6, 35.8, 45, 54, 64.2, 73.7, and 83.5 MJ when offsets of 60, 50, 40, 30, 20, 10, 5 m, and zero are employed, respectively.

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Figure 6 – Total energy release (TER) when retreating longwalls in the lower seam in the San Hejian mine The Journal of The Southern African Institute of Mining and Metallurgy

Evaluating the coal bump potential for gateroad design in multiple-seam longwall mining

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Figure 7 â&#x20AC;&#x201C; Total energy release (TER) when retreating panel 9113 with different offsets

Evaluating the coal bump potential for gateroad design in multiple-seam longwall mining FORREST, P., HAYCOCKS, C., and ZHOU, Y. 1987. Design of lower seam longwall operations in multiple seam mining. Proceedings of the 6th International Conference on Ground Control in Mining, West Virginia University, Morgantown, WV. Peng, S.S. (ed.).

HAYCOCKS, C. and KARMIS, M. 1983. Ground Control Mechanisms in Multi-Seam Mining. Final report submitted to US Bureau of Mines, Office of Mineral Institutes.

HAYCOCKS, C. and ZHOU, Y.X. 1990. Multiple-seam mining – a state of-the-art review. Proceedings of the 9th International Conference on Ground Control in Mining, West Virginia University, Morgantown, WV. Peng, S.S. (ed.).

HEASLEY, K.A. 1998, Numerical Modeling of Coal Mines with a Laminated Displacement-Discontinuity Code. PhD dissertation, Colorado School of Mines.

Figure 8 – Peak TERs vs offset distances in panel 9113

HEASLEY, K.A., WORLEY, P., and ZHANG. Y. 2003. Stress analysis and support design for longwall mine-through entries (a case study). Proceedings of the 22nd International Conference on Ground Control in Mining, West Virginia University, Morgantown, WV. HEASLEY, K.A. and AGIOUSTANTIS, Z.G. 2007. LaModel: a boundary-element program for coal mine design. Proceedings: New Technology for Ground Control in Multiple-seam Mining. IC 9495. National Institute for Occupational Safety and Health, Washington, DC. HOLLAND, C.T. 1973. Mine pillar design. SME Mining Engineering Handbook, vol. 1, section 13.8. AIME. KARABIN, G. and EVANTO, M. 1994. Experience with the boundary-element method of numerical modeling to resolve complex ground control problems. Proceedings of the 18th International Conference on Ground Control in Mining, West Virginia University, Morgantown, WV.

Figure 9 – Measured entry convergences during panel 9113 retreat

bumps according the empirical criteria proposed in this paper.

Hence, to be more conservative, an offset of 60 m was ultimately adopted when developing the gateroads of panel 9113 in the field. Field observation showed that coal bumps were avoided successfully and the maximum roof-to-floor and rib-to-rib convergences were only 360 mm and 576 mm, respectively.

In this study, the use of TER for evaluation of coal bump potential was validated, based on past experience of successful working. However, the method was unsuccessful regarding coal bump potential in Crandall Canyon Mine in Utah (Sears and Heasley, 2009; Sears, 2009).Notwithstanding, the use of TER for coal bump risk assessments shows promise.

Acknowledgments This work was supported by the Self-research Program of the State Key Laboratory of Coal Resources and Safety Mining of China through contract no. SKLCRSM08X04, the National Natural Science Foundation of China through contracts no.

51204166 and 51174195, and the Priority Academic Program Development of Jiangsu Higher Education Institutions through contract no. SZBF2011-6-B35. This support is gratefully acknowledged. The authors also wish to thank Dr. Keith Heasley (West Virginia University) for his valuable comments on this paper during its preparation.

References CHEKAN, G.J. and LISTAK, J.M. 1993. Design Practices for Multiple-Seam Longwall Mines. IC 9360. US Bureau of Mines.

DOU, L. and HE, X. 2001. Theory and Technology of Rock Burst Prevention.

China University of Mining and Technology Press, Xuzhou.

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LI, H., LIU, M., KANG, Q., and ZHAI, X. 2000. Optimization of district mine layout in multi-seam mining: a case study. Proceedings of the 19th International Conference on Ground Control in Mining, West Virginia University, Morgantown, WV. Peng, S.S. and Mark, C. (eds).

MALEKI, H.N., STEWART, C., and STONE, R. 2007. Three-seam stress analyses at Bowie Mines, Colorado. Proceedings of the 26th International Conference on Ground Control in Mining, West Virginia University, Morgantown, WV. Peng, S.S. et al. (eds).

MATETIC, R.J., CHEKAN, G.J., and GALEK, J.A. 1987a. Design considerations for multiple-seam mining with case studies of subsidence and pillar load transfer. Proceedings of the 28th US Symposium on Rock Mechanics, University of Arizona, Tucson, AZ. MATETIC, R.J., CHEKAN, G.J., and GALEK, J.A., 1987b. Pillar Load Transfer Associated with Multiple-Seam Mining. RI 9066. US Bureau of Mines.

MARK, C. 2007. Multiple-seam mining in the Central Appalachian coalfields.

Proceedings of the New Technology for Ground Control in Multiple-Seam Mining. IC 9495. National Institute for Occupational Safety and Health, Washington, DC. ORAM, J.S. and PONDER, C.G. 1997. Measurement of effects of interaction and influence on mine layout design at Maltby Colliery. Proceedings of the 16th International Conference on Ground Control in Mining, West Virginia University, Morgantown, WV. Peng, S.S. (ed.).

PENG, S.S. and CHANDRA, U. 1980. Getting the most from multiple seam reserves. Coal Mining and Processing, vol. 17. pp. 78–84.

PENG, S.S. 2007. Ground Control Failures – a Pictorial View of Case Studies.

University pf West Virginia, Morgantown, W.V. PENG, S.S. 2008. Coal Mine Ground Control. 3rd edn. University pf West Virginia, Morgantown, WV. SALAMON, M.D.G. 1990. Mechanism of caving in longwall coal mining. Rock Mechanics Contribution and Challenges. Proceedings of the 31st US Symposium on Rock Mechanics, Golden, CO. CRC Press, Boca Raton, FL. SEARS, M. and HEASLEY, K.A., 2009. An application of energy release rate.

Proceedings of the 28th International Conference on Ground Control in Mining, West Virginia University, Morgantown, WV. Peng, S.S. (ed.).

SEARS, M. 2009. Implementing Energy Release Rate Calculations into the LaModel Program. MS thesis, West Virginia University, Morgantown, WV. STEWART, C., STONE, R., and HEASLEY, K.A., 2006. Mine stability mapping.

Proceedings of the 25th International Conference on Ground Control in Mining, West Virginia University, Morgantown, WV. ZIPF, R.K. 2005. Failure mechanics of multiple seam mining interaction.

Proceedings of the 24th International Conference on Ground Control in Mining, West Virginia University, Morgantown, WV. Peng. S.S. (ed.) ◆

The Journal of The Southern African Institute of Mining and Metallurgy

http://dx.doi.org/10.17159/2411-9717/2015/v115n8a13

Influence of medium particle size on the separation performance of an air dense medium fluidized bed separator for coal cleaning

by S. Mohanta* and B.C. Meikap†‡

Dry beneficiation of coal by air dense medium fluidized bed is an emerging trend. It is widely believed that the particle size of the medium has a significant effect on the separation efficiency. This investigation demonstrates that medium particle size has a major effect on separation efficiency. Experimental results show that different size fractions of the same feed respond differently to the same size fraction of medium solids.

Furthermore, a particular size fraction of feed coal responds differently with different size fractions of medium solids. The Ep values and overall metallurgical performance parameters, obtained from experimental results, clearly indicate the superior performance of an air dense medium fluidized bed separator when using –150+106 μm magnetite powder as the fluidizing medium. These observations reinforce the importance of sizing the medium particle size for the air dense medium fluidized bed.

Keywords coal beneficiation, air dense medium fluidized bed, separation efficiency, partition curve.

Introduction In an air dense medium fluidized bed (ADMFB), a fluidizing bed of a particular density is created by suspending solid particles in an upward air flow. A feed containing coal particles of different densities is introduced into this fluidizing bed. The particles with relative density less than the bed density float on the top of the bed, while particles denser than the bed density sink through it. Since there is a general correlation between ash content and density of coal particles, it is possible to remove ash-forming impurities from a raw coal to the required level by regulating the density of the fluidizing bed. Recently, the ADMFB process has received significant attention world-wide for its specific advantages of low operating cost and less restrictive for small capacities (Luo and Chen, 2001; Houwelingen and de Jong, 2004; Dwari and Hanumantha Rao, 2007; Sahu et al., 2009; Mohanta et al., 2012). This process is not recognized as a well-established industrial practice because of the availability of efficient wet processing technology. However, with the revival of interest in dry beneficiation of coal in general, and ADMFB in particular, there is a concerted The Journal of The Southern African Institute of Mining and Metallurgy

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* Department of Chemical Engineering, Indira Gandhi Institute of Technology, Sarang, India. † Department of Chemical Engineering, Indian Institute of Technology, Kharagpur, India. ‡ School of Chemical Engineering, University of KwaZulu-Natal, Durban, South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Aug. 2014 and revised paper received Oct. 2014 AUGUST 2015

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Synopsis

effort to improve the efficiency of the ADMFB process, and successive improvements have made the process a competitive and viable option for beneficiating coal.

It is widely believed that efficient beneficiation of coal in an ADMFB separator can be achieved only if the fluidized bed is a microbubble fluidized bed with a uniform and stable density, as well as with minimal backmixing of the medium solids (Akiyama and Yoshikawa, 1999). However, several researchers (Yasui and Johanson, 1958; Baumgarten and Pigford, 1960; Rowe and Stapleton, 1961) have claimed that the bubble size increases with increased medium particle size, and that the formation of microbubbles in beds of large particles is difficult. On the other hand, uniform and stable fluidization of smaller particles is difficult because the interparticle forces are greater than the force that the fluid can exert on the particles. It is also realized that the size of the fluidizing medium plays an important role in the separation performance of an ADMFB separator. Therefore, in order to improve the separation efficiency, accurate determination of the appropriate size range of medium particles is of paramount importance.

Many researchers have studied the kinetics of the fluidization process and used different size ranges of magnetite powder; however, most of them (Luo and Chen, 2001; He et al., 2002; Choung et al., 2006; Zhao et al., 2010) suggest that the size range of 150 to 300 μm is the most suitable for efficient

Influence of medium particle size on the separation performance beneficiation of coal by overlooking the size of coal to be processed. In actual practice, the mixing of medium solids in an ADMFB is caused by the disturbance created by the air bubbles passing through the bed. Moreover, the rate of growth of disturbance depends primarily on the size of the fluidizing medium solids. Therefore, to improve the separation efficiency, the use of medium or small size particles would be beneficial as it allows for a low superficial air velocity that is sufficient to minimize back-mixing of the medium solids. However, reducing the size of the medium solids increases the viscosity of the bed, thus restricting the separation of the coal particles. Furthermore, during beneficiation fine coal particles are accumulated in the bed and alter the fluidization behaviour of the magnetite particles.

Thus, during the beneficiation of coal, the appropriate size range of medium particles is required in conjunction with optimum superficial gas velocity to achieve full separation efficiency. This investigation aims to demonstrate that to improve the separation efficiency of an ADMFB in general, a distinct size fraction of medium is needed so that an ideal superficial gas velocity can be established. An ideal superficial gas velocity should be sufficient to create a stable pseudo-fluid medium bed, while low enough to avoid backmixing.

Experimental A schematic representation of the fluidized bed used in this study is shown in Figure 1. The fluidizing vessel is a Perspex vertical cylindrical column with 15 cm inside diameter and 150 cm height. A conical chamber 42 cm in height, situated at the bottom of the vessel and connected to an air inlet pipe, acts to create a uniform distribution of air throughout the entire cross-section of the fluidizing column.

To prevent a direct air jet impinging onto the distributor and to ensure uniform fluidization, a dissipative structure is built

onto the inlet air pipe. To generate the microbubbles, multiple layers of filter cloth with an aperture size of 3–5 μm and 4.9% air permeability, supported on a wire mesh, are used as the distributor plate. A ruler is fixed along the length of the column to measure the fluidized bed height. A series of manometers are connected at different heights around the fluidization column to measure the pressure drop across the bed and to check the uniformity and stability of the bed. A bag filter is connected at the top of the fluidization column to collect the entrained dust particles from the outlet air.

The magnetite powder used in this investigation was prepared from a sample obtained from Bakla mines of Odisha, India. The particle size was reduced by a Blake jaw crusher, followed by roll crushing and ball milling. After size reduction, the magnetite was separated from nonmagnetic impurities using a hand magnet. The material was then screened to different size fractions and the particle size distribution of each fraction determined by Malvern particle size analyser (Model 2000E).

The fluidization characteristics of magnetite powder depend mostly on its size and size distribution. In general, fine magnetite particles tend to clump and agglomerate if their moisture content is above 4%. Moreover, the fine particles in a wide size distribution can be fluidized with a wide range of gas flow rates. On the other hand, beds of larger magnetite particles fluidize poorly, with bumping, spouting, and slugging. Therefore, magnetite particles of a particular size with narrow size distribution are required in an ADMFB for coal beneficiation. For a uniform and stable fluidized bed, the acceptable magnetite particle size range is selected from a Geldart fluidization classification. In order to achieve an effective separation, the magnetite particles and coal particles should be of different Geldart groups, and the magnetite particles need to be more readily fluidized than the coal particles.

Figure 1 – Schematic diagram of the experimental set-up. 1. Air compressor; 2. Surge tank; 3, 5. Pressure gauge; 4. Pressure regulator; 6. Solenoid valve; 7, 10. Needle valve; 8. Gate valve; 9. Rotameters; 11. Dissipative structure; 12. Distributor plate; 13. Fluidized bed; 14. U-tube manometers; 15. Bag filter

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The Journal of The Southern African Institute of Mining and Metallurgy

Influence of medium particle size on the separation performance Results and discussion

The effect of medium size on separation performance of the ADMFB separator was studied by considering five different size fractions of magnetite particles: −300+212, −150+106, −106+75, −75+63, and −63.0 μm. The size distributions of these fractions are shown in Figure 2. Two different size fractions of coal (−50+25 and−25+13 mm) obtained from Jagannath opencast mines of Odisha, India were considered for this investigation. To ascertain the correct magnetite size, coal samples with different characteristics were tested with the selected magnetite size.

All the experiments were conducted at room temperature.

The magnetite particles were loaded on the air distributor to a height of 22 cm. Compressed air was supplied to the ADMFB column to fluidize the magnetite powder, and the air flow rate was adjusted to maintain the required bed height to achieve the desired bed density for separation. The air flow rate through the air distributor and the supplied air pressure were kept constant at 3.5 × 10−3 m3/s and 49 kPa, respectively. About fifteen minutes was allowed for the fluidized bed to stabilize. A weighted amount of feed coal was then slowly introduced into the top surface of the fluidized bed. After stratification for 30 seconds, the inlet air supply was suddenly shut off so that all the stratified/segregated particles were retained in their original position in the defluidized bed. The mixture of medium particles and coal particles in the separator was carefully collected into two fractions, from the top to the bottom, as product and reject respectively.

The coal particles from each test were separated from the medium particles by screening and subjected to float-sink analysis using a range of equally spaced density intervals from 1.2 to 2.0. Ultimate analysis of the coal in each density class was conducted by the recommended procedure IS: 1350 (Part I) (Bureau of Indian Standards, 1984). The yield of clean coal was determined by direct mass measurement. The float-sink data obtained from the clean coal and the reject samples were used to compute the so-called ‘reconstituted feed’. The reconstituted feed reflects the washability analysis of the coal fed into the ADMFB unit.

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Figure 3 – Partition curves generated for different size magnetite powders for Jagannath coal (a) -50+25 mm, (b) -25+13 mm VOLUME 115

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Figure 2 – Particle size distributions for magnetite powders

Both the collected coal samples were washed separately in the ADMFB using five different sizes of magnetite powder as fluidizing medium. As expected, the misplacement of lowdensity clean coal in high-density reject coal increases as the medium particle size decreases. More importantly, when the fluidizing bed consisted of −300+212 μm size magnetite powder, all the coal particles remained at the top surface of the fluidized bed, whereas with −63 μm size magnetite powder all the coal particles remained at the bottom surface of the bed. It is clear from the ADMFB separation test that a lower superficial gas velocity is required to fluidize the finer sized medium particles. At such a lower gas velocity, the upward drag of the rising gas bubbles is not sufficient to sweep the descending light coal particles back to the top surface of the bed. Therefore, the lighter clean coal particles descend into the bed and fall to the bottom, remaining there together with the heavy coal particles. The situation is quite the opposite for the −300+212 μm medium particles. The required higher superficial gas velocity exerts a lifting effort great enough to push the heavy coal particles towards the

Influence of medium particle size on the separation performance top of the bed, thus all the coal particles remain at the top surface of the bed.

It is also expected that different size fractions of the same feed coal will respond differently to a particular size fraction of medium solids, which may affect the separation efficiency of the ADMFB. To study this aspect, partition curves were generated separately from the test results for both size fractions of feed coal. Figure 3 shows that the same size fraction of feed coal responds differently to different size fractions of medium solids. Also, the different size fractions of a given feed respond differently with the same size fraction of medium solids. The Ecart Probable Moyen (Ep) (Mohanta et al., 2011) values of −50+25 mm size coal for −75+63, −106+75, and −150+106 μm size medium solids are calculated as 0.09, 0.04, and 0.01, respectively. Similarly, the calculated Ep values of −25+13 mm size coal for −106+75 and −150+106 μm size medium solids are 0.17 and 0.08, respectively. The lower the Ep value, the nearer to vertical is the partition curve between 25% and 75% and the more efficient the separation. It can be clearly seen from Figure 3, as well as from the Ep values, that magnetite particles in the size range −150+106 μm are the most suitable for beneficiation of both −50+25 and −25+13 mm size coals. However, for the same size fraction of medium solid the misplacement of clean coal in reject coal is greater for smaller size fraction coals. The apparent discrepancy is attributed to multibody interactions between bed particles and feed materials and to instability of the medium created by disturbing bubbles. Such disturbances are more severe in case of very fine medium solids.

To ascertain the correct size range of medium solids, −50+25, and −25+13 mm size Jagannath coal and −25+13 mm size Hingula coal were washed in an ADMFB separator using −150+106 μm size magnetite powder as the medium solid. Fluidized beds of different densities were formed by mixing −150+106 μm magnetite powder and −300+212 μm coal powder in different proportions and maintaining appropriate experimental conditions. All the coals were washed at different bed densities, and the partition curves and separation efficiencies were calculated from the test results.

Figure 4 shows the partition curves generated for the three coal samples. It can be seen that excellent separation performance is achieved with the −150+106 μm size magnetite powder, and the misplacements of the clean and reject coals are minimal. Also, the separation performance at lower specific gravity of separation/cut-point (d50 = 1.62) is quite satisfactory, with an Ep value of 0.11. To assess the basic shape compatibility, it is important to compute and display the partition curve over a range extending well beyond both the upper and lower limits of the specific gravity of separation. Furthermore, sometimes even a small specific gravity error will cause a large equivalent partition error when the partition curve is very steep, as in the case of the partition curve for −50+25 mm size coal (Figure 4a).

Therefore, to encompass a broad spectrum of distribution curves generated at different specific gravities of separation

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and to facilitate comparison between them, generalized partition curves are generated (Figure 5). A conventional partition curve can be obtained from this generalized curve simply by multiplying the reduced specific gravity (the abscissa) by the specific gravity of separation. An equation of the following form provides an extremely good fit for the generated generalized distribution curve data and may be used for simulation programs.

[1]

where y is the percentage distribution of clean coal corresponding to x; x is the normalized specific gravity; a, b, and c are parameters in the function. Parameters b and c are the low and high specific gravity limits of the partition function respectively, the parameter a is very closely related to sharpness of separation. Furthermore, 95% confidence limits are placed on the equation parameters, which gives an indication of the error involved in estimating them.

To compare the different medium size fractions, different metallurgical performance parameters are calculated for each size fraction of coal (Table I). The performance parameters can be defined as

[2]

[3]

[4]

Separation efficiency (%) = Combustible recovery – Ash recovery

[5]

where Yc is the clean coal yield (%), and Ac and Af are ash contents (%) of clean coal and feed, respectively.

The tabulated organic efficiency values indicate that the misplacement of the clean coal and the reject coal is minimum while using −150+106 μm size magnetite powder as fluidizing medium. The values of combustible recovery are quite high for this medium size, but the separation efficiency is significantly low due to the higher yield and thus lower ash rejection. The overall metallurgical performance parameters point towards the superior performance of the ADMFB separator while using −150+106 μm size magnetite powder as fluidizing medium.

The Journal of The Southern African Institute of Mining and Metallurgy

Influence of medium particle size on the separation performance

The Journal of The Southern African Institute of Mining and Metallurgy

Figure 5 – Generalized distribution curves for different coals. (a) −50+25 mm Jagannath coal, (b) −25+13 mm Jagannath coal, (c) −25+13 mm Hingula coal VOLUME 115

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Figure 4 – Partition curves for different coals. (a) −50+25 mm Jagannath coal, (b) −25+13 mm Jagannath coal, (c) −25+13 mm Hingula coal

Influence of medium particle size on the separation performance Table I

Performance of ADMFB separator for different medium sizes Coal

Size of medium solid (µm)

Combustible recovery (%)

Ash rejection (%)

Separation efficiency (%)

Organic efficiency (%)

Jagannath coal

-150+106

95.39

16.10

11.50

99.53

(-50+25 mm)

-106+75

92.45

22.33

14.78

98.21

-75+63

52.58

67.05

19.63

73.96

Jagannath coal

-150+106

89.86

27.74

17.59

97.05

(-25+13 mm)

-106+75

52.54

66.33

18.87

67.44

-75+63

33.31

79.62

12.93

44.79

Conclusions

DWARI, R.K. and HANUMANTHA RAO K. 2007. Dry beneficiation of coal – A

In this study the effect of size of the fluidizing medium solid on the separator performance of an ADMFB was investigated, and an attempt made to determine the optimum size range over which the separator can operate satisfactorily. The results clearly show that different size fractions of the same feed coal respond differently with the same size fraction of medium solid and vice versa.The misplacement of low-density clean coal in high-density reject coal increases as the medium particle size decreases, thus reducing the separation performance. On consideration of the Ep values and different metallurgical performance parameters, −150+106 μm size magnetite powder is found to be the most suitable for washing of coal in the size range of −50+13 mm.

However, it is important to note that this best medium size is subject to the particular design of the distributor plate.

It appears that the air dense medium fluidized bed method is not suitable for beneficiation of coal of -13+1 mm size, since the separation efficiency decreases sharply at coal particle sizes less than 13 mm and the process become uneconomical. Wet separators such as dense medium cyclones and water-only cyclones are more effective and economical over this smaller size range.

The design of a continuous air dense medium fluidized bed of rectangular cross-section with 1500 mm length, 1200 mm height, and 108 mm width with 600 kg/h capacity is currently in progress, and we have applied for funding for a pilot plant study.

review. Mineral Processing and Extractive Metallurgy Review, vol. 28. pp. 177–234.

HE, Y., ZHAO,Y., CHEN, Q., LUO, Z., and YANG, Y. 2002. Development of the density distribution model in a gas-solid phase fluidized bed for dry coal separation. Journal of the South African Institute of Mining and Metallurgy, vol. 102, no. 7. pp. 429–434.

HOUWELINGEN, J.A. and VAN AND DE JONG, T.P.R. 2004. Dry cleaning of coal: review, fundamentals and opportunities. Geologica Belgica, vol. 7. pp.

335–343.

LUO, Z. and CHEN, Q. 2001. Dry beneficiation technology of coal with an air dense-medium fluidized bed. International Journal of Mineral Processing, vol. 63. pp. 167–175.

MOHANTA, S., DARAM, A.B., CHAKRABORTY, S., and MEIKAP, B.C. 2012.

Applicability of the air dense medium fluidized bed separator for cleaning of high-ash Indian thermal coals: an experimental study. South African Journal of Chemical Engineering, vol. 16. pp. 50–62.

MOHANTA, S., CHAKRABORTY, S., and MEIKAP, B.C. 2011. Influence of coal feed size on the performance of air dense medium fluidized bed separator used for coal beneficiation. Industrial and Engineering Chemistry Research, vol. 50. pp. 10865–10871.

ROWE, P.N. and STAPLETON, W.M. 1961. The behaviour of 12-inch diameter gas fluidized beds. Transactions of the Institution of Chemical Engineers, vol.

39. pp. 181–187.

SAHU, A.K., BISWAL, S.K., and PARIDA, A. 2009. Development of air dense

References

medium fluidized bed technology for dry beneficiation of coal – a review.

AKIYAMA, T. and YOSHIKAWA, T. 1999. Effects of vessel material on the air pressure distribution within vibrating particle beds. Powder Technology,

International Journal of Coal Preparation and Utilization, vol. 29. pp. 216–241.

vol. 103. pp. 139–144.

BAUMGARTEN, P.K. and PIGFORD, R.L. 1960. Density fluctuations in fluidized beds. AIChE Journal, vol. 6. pp. 115–123.

BUREAU OF INDIAN STANDARDS. 1984. IS: 1350, (Part I). Methods of test for coal and coke (reaffirmed 2001). New Delhi.

dense medium fluidized bed. Coal Preparation, vol. 26. pp. 1-15.

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beds. AIChE Journal, vol. 4. pp. 445–452.

ZHAO, Y., TANG, L., LUO, Z., LIANG, C., XING, H., WU, W., and DUAN, C. 2010.

Experimental and numerical simulation studies of the fluidization charac-

CHOUNG, J., MAK, C., and XU, Z. 2006. Fine coal beneficiation using an air

766

YASUI, G. and JOHANSON, L. N. 1958. Characteristics of gas pockets in fluidized

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teristics of a separating gas-solid fluidized bed. Fuel Processing Technology, vol. 91. pp. 1819–1825.

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Measuring and modelling of density for selected CaO-Al2O3-MgO slags

by J.F. Xu*, K. Wan*, J.Y. Zhang†, Y. Chen*, and M.Q. Sheng*

The densities of selected CaO-Al2O3-MgO slag systems were measured at 1823K by the Archimedean method. Thirteen different slag compositions were chosen based on different levels of the MgO content and the mass ratio of CaO/Al2O3. The results indicated that the density of the slag decreases with increasing MgO content (from 0–3.78 mass%), but increases with further increases in MgO content up to 11.33 mass%. At a fixed MgO content of 5.5%, the trend of density change with CaO/Al2O3 in the slag is similar to that for changes in MgO content. On the basis of the regular solution approximation rules of excess molar quantities, an attempt was made to estimate the molar volume of the slags investigated.

The application of the molar volume model confirmed that the present expanded approximation rules are applicable to predict the molar volumes of the melts discussed.

Keywords slag density, CaO-Al2O3-MgO system, molar volume, modelling.

Introduction Density is one of the most important fundamental properties of slags, which influences metallurgicaly phenomena in many ways. Density is sensitive to the nature of the chemical bonds and the species in the melt, and is of great importance not only for theoretical research on the structural properties of molten slags, but also for industrial applications. Density is also required to estimate other key properties used to assess the behaviour of high-temperature molten oxides, e.g. viscosity, surface tension, and thermal conductivity.

Liquid calcium aluminate slags that contain magnesia are the basis of most ladle slags for secondary refining processes. The CaO-Al2O3 binary phase diagram (Hallstedt , 1990) shows the eutectic composition (50 wt% CaO and wt% Al2O3) has the lowest melting temperature. The main emphasis of this study is on the 12CaO·7Al2O3 refining slag with MgO additions. The CaO-Al2O3-MgO ternary system is of considerable importance in industrial metallurgical processes, and is fundamental to the understanding of metallurgical slags, ceramic materials, and geological phenomena.

Many researchers have reported on the optimization of the ternary CaO-Al2O3-MgO system (Hallstedt, 1995; Jung, Degterov, and The Journal of The Southern African Institute of Mining and Metallurgy

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Experimental Sample preparation and characterization The CaO-Al2O3-MgO equilibrium phase diagram from Verlag Stahleisen (Slag Atlas, 1995) is shown in Figure 1. Table Ι shows the nominal chemical compositions of samples used in the present work. The samples were divided into two groups. In the first group,

* Shagang School of Iron and Steel, Soochow University, China. † Shanghai Key Lab. of Modern Metallurgy and Material Processing, Shanghai University, China. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Oct. 2013 and revised paper received Feb. 2015.

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Synopsis

Pelton, 2004). It has been shown that the liquidus temperature of the slag is lower than 1823K when the addition of MgO to 12CaO·7Al2O3 is less than 10 wt%. The density of the CaO-Al2O3-MgO system has not been systematically investigated experimentally, and few experimental studies of density have been reported in the CaO-Al2O3MgO system (Slag Atlas, 1995). Due to the difficulty of experimental measurements at high temperature, the existing data covers only a limited composition and temperature range.

In our previous work, the viscosity of the selected CaO-Al2O3-MgO slag system was measured by using the rotating cylinder method and the effects of temperature, the MgO content, and the mass ratio of CaO/Al2O3 were studied. The results indicated that both the MgO content and the mass ratio of CaO/Al2O3 influence the viscosity of the slag (Xu et al., 2011). The aim of the present work is to investigate the effect of MgO additions on the density of 12CaO·7Al2O3 slag. The effects of the MgO content and the mass ratio of CaO/Al2O3 were studied. On the basis of the regular solution approximation rules of excess molar quantities, an attempt was made to estimate the molar volumes of selected CaOAl2O3-MgO slags investigated.

Measuring and modelling of density for selected CaO-Al2O3-MgO slags

Figure 1 – Liquidus projection of CaO-Al2O3-MgO slag system (Slag Atlas)

comprising five samples, the mass ratios of CaO/Al2O3 were equal to unity, while the MgO content was varied from zero to 12 wt%. The eight samples in the second group had MgO contents of 5.5 wt%, and the mass ratio of CaO/Al2O3 was varied from 0.60 to 1.20.

Samples were prepared from CP (chemically pure) grade CaO (≥99.0 wt%), Al2O3 (≥99.5 wt%), and MgO (≥99.0 wt%) powders, which were dried at 373K for 24 hours. The method for preparation of the slag samples has been reported in detail elsewhere (Xu et al., 2012). The powders were ground and

weighed to the desired compositions and mixed in a mortar, and the mixtures were melted in a graphite crucible in an air induction furnace for 30 minutes at 1773K. The fused slag samples were poured onto the surface of a cold steel plate. For further homogenization, these samples were then crushed and ground to fine powders. The powder samples were placed in a corundum crucible and were dried and decarburized at 1223K for 30 hours in a muffle furnace in air. Finally, the chemical compositions of the samples were analysed; the results are reported in Table Ι.

Table Ι

Nominal and analysed chemical compositions of slag samples, wt% Sample no.

Nominal CaO

1-1 50.00 1-2 48.00 1-3 47.00 1-4 46.00 1-5 44.00 2-1 47.00 2-2 46.00 2-3 44.00 2-4 42.00 2-5 40.00 2-6 35.44 2-7 49.50 2-8 51.54 * R is the mass ratio of CaO/Al2O3

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Density

Al2O3

MgO

R*

CaO

Al2O3

MgO

R

g/cm3

50.00 48.00 47.00 46.00 44.00 47.50 48.50 50.50 52.50 54.50 59.06 45.00 42.96

0.00 4.00 6.00 8.00 12.00 5.50 5.50 5.50 5.50 5.50 5.50 5.50 5.50

1.00 1.00 1.00 1.00 1.00 0.99 0.95 0.87 0.80 0.73 0.60 1.10 1.20

49.80 48.02 46.61 46.26 44.16 46.62 47.08 44.29 42.83 41.00 35.25 51.14 52.40

49.46 48.43 46.90 45.58 43.84 47.53 47.96 50.26 52.36 53.54 58.60 43.58 40.94

0.39 3.78 5.79 7.60 11.33 5.24 4.86 5.05 4.96 4.91 5.14 5.50 5.24

1.01 0.99 0.99 1.01 1.01 0.98 0.98 0.88 0.82 0.77 0.60 1.17 1.28

2.70 2.47 2.72 2.78 3.51 3.22 2.70 2.75 3.14 2.93 3.50 3.32 3.49

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The Journal of The Southern African Institute of Mining and Metallurgy

Measuring and modelling of density for selected CaO-Al2O3-MgO slags

Figure 4 – Relationship between density and MgO content Figure 2 – Schematic of experimental apparatus

Figure 5 – Relationship between density and ratio of CaO/Al2O3

Density measurements The densities were measured at 1823K by the Archimedean method using the RTW-08 type testing instrument . The experimental set-up and procedure have been described in detail in an earlier publication (Xu et al., 2012). In this section, a brief description of the method is given. The furnace (Figure 2) had a maximum temperature of 1873K. The temperature was measured by a Pt-30Rh/Pt-6Rh thermocouple touching the crucible bottom from outside. The metal bob for measurement was made of molybdenum, which has a melting point of 2896K. The volume of the bob at high temperature was calculated from the values measured in pure water in the temperature range 283–308K and the coefficient of thermal expansion of molybdenum. Dimensions of the crucible and bob are presented in Figure 3. The size of bob had to be carefully designed so that it was applicable in the required measurement range. Purified argon gas (0.2 L/min) was introduced into the reaction chamber during the entire process. The graphite crucible was filled with 120 g slag and placed in the furnace, and the furnace was programmed to heat up to 1823K at a heating rate of 10K/min. The furnace was kept at the target temperature at least for 30 minutes.

The densities of the slags were then measured at the The Journal of The Southern African Institute of Mining and Metallurgy

predetermined temperature, and the measurements were repeated several times until the values were stable, at which the error was about ±1×10-3 g. After measurement was complete, the furnace was allowed to cool.

In the density measurements several sources of error may occur. As mentioned above, a deviation in the volume of the bob caused an error of ±0.5% in density, the determination of the temperature with an accuracy of ±5K introduced an error of ±0.3% in density, and the accuracy of ±1×10-3 g in the weighting balance caused an additional error of ±0.3% in density. Since the effect of surface tension on the density measurement is difficult to estimate, no corrections were made for the effect of surface tension of the melt acting on the thin section of the spindle. This has been calculated to cause an error of about 2.0% in the density measurement (Nakanishi et al., 1998). The total error in the determination of the density was less than ±3.1%.

Results and discussion The density of the selected CaO-Al2O3-MgO slag system was measured at 1823K. Thirteen different slag compositions were chosen based on different levels of MgO content and mass ratios of CaO/Al2O3. The MgO content was varied from 0.39 to 11.33 wt.%, and CaO/Al2O3 mass ratios varied between 0.60 and 1.28. The effects of MgO content and CaO/Al2O3 ratio on the density are shown Figure 4 and Figure 5 respectively.

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Figure 3 – Dimensions of crucible and bob

Measuring and modelling of density for selected CaO-Al2O3-MgO slags The data for corresponding slags from the literature (Slag Atlas, 1995) is also shown in Figures 4 and Figure 5.

Comparison reveals that although the values obtained in the present study may be higher or lower than those from the literature, the density trends are similar. The differences in densities could be related to slag composition and experimental design and procedure. It should be noted that there are difficulties in retrieving reliable experimental density values.

Effect of MgO content on density The measured densities at a constant mass ratio of CaO/Al2O3 and various amounts of MgO (0.39–11.33%) at 1823K are shown in Figure 4. The results indicate that the density of the slag decreased for low additions of MgO, with a minimum value of 2.47g/cm3 at 1823K at an MgO content of 3.78%.

This was followed by a sharp increase in density with further additions of MgO. An additive method for the estimation of density in alloys and slags has been widely used for some time. The densities of the pure components CaO, MgO, and Al2O3 are 3.32, 3.50, and 3.97g/cm3 respectively (Lide, 2003). Since the density of MgO is much lower than that of Al2O3, slag density will decrease with small increases in MgO content. On the other hand, the addition of metal oxides has a strong impact on the physical properties of molten slag systems. The effect of cations on the structure is commonly related to the charge of the cations as well as the radius. In order to estimate the effect of different cations, the ratios z/r or z/r2 are used, where z is the valence and r is the cation radius. Slags are composed of cations and complex anions, and the forces of attraction between ions directly affect slag density (Cui et al., 1996). When MgO is added to 12CaO·7Al2O3-type slags, the complex polymers of aluminum such as AlO45- tetrahedra break down into smaller units, decreasing the degree of polymerization, and the radius of the ions decreases; then the force of attraction between ions and the density of the slag increase (Cui et al., 1996; Mills, 1993). Furthermore, the molar masses of CaO, MgO, and Al2O3 are 56.08, 40.31, and 101.96 g/mol, respectively. Since the molar mass of MgO is lower than that of the other metal oxide components of the slags, when MgO replaces part of the CaO and Al2O3 content in 12CaO·7Al2O3, the total number of ions in the molten slag will increase. This also increases the forces of attraction between the ions, and thus increases the density of the slag.

increasing the attractive forces between ions. Furthermore, the total number of ions in the molten slag will also increase with increasing of the mass ratio of CaO/Al2O3. The lower degree of polymerization and the higher number of ions will enhance the attractive forces between ions, and the density of the slag will increase (Lide, 2003). Due to these effects, the density of slag at first decreases with increasing the mass ratio of CaO/Al2O3, followed by an increase in density with further increases in the mass ratio of CaO/Al2O3. The lower density values of slag at CaO/Al2O3 mass ratios in the range 0.9–1.0 could also be due to the fact that the slag compositions are between the two eutectic points in the CaOAl2O3-MgO slag system, as shown in Figure 1.

Estimated molar volume of the selected slags The molar volume, which the reciprocal of the density multiplied by the molar mass, is an important thermodynamic property. Due to the inherent difficulties associated with measurements at high temperature, it is necessary to have access to reliable models for estimating the molar volume of slags, which are reflective of the structure of the melt. There are many kinds of model cited in the literature (Zhang and Chou, 2010; Persson, Matsushita, and Zhang, 2007; Bottinga, Weill, and Richet, 1982; Mills, Yuan, and Jones, 2011; Hayashi, Abas, and Seetharaman, 2004; Priven, 2004; Nakajima, 1994; Vadasz, Havlik, and Danek, 2006; Shu, 2007; Zhang and Chou, 2009) for estimating the molar volume of slag, including physical models and semi-empirical models. Physical models, which are based on the structure of atoms and molecules, can give a clear physical picture of the practical solution. The semi-empirical models combine both theoretical considerations and practical thermodynamics; these models can give more reasonable data and be suitable for many systems with larger compositional ranges.

In order to estimate molar volumes for multi-component silicate melts, expanded approximation rules are proposed, on the basis of the regular solution approximation rules of excess molar quantities for a binary system melt (Vadasz, Havlik, and Danek, 2006). A brief description of the model is given below. Detailed discussions about this method can be found in Vadasz, Havlik, and Danek, (2006).

The molar volume of slags is calculated from Equation [1]: [1]

Effect of CaO/Al2O3 on density The effect of the mass ratio of CaO/Al2O3 (in the range 0.6–1.28) on density at different temperatures and a constant MgO content of 5.5 wt.% is shown in Figure 5. The effect of CaO/Al2O3 mass ratio on density was the same as the effect of MgO content; density decreases at first, and then increases with increasing the mass ratio of CaO/Al2O3. The minimum density at 1823K was 2.70g/cm3 at a CaO/Al2O3 mass ratio of 0.98.

Since the density of pure CaO is lower than that of Al2O3, increasing CaO/Al2O3 decreases the slag density in line with the additive method of density calculation. On the other hand, with increasing mass ratio of CaO/Al2O3, the networkbreaking cations (Ca2+) present in slag increase, and the complex polymers of aluminum such as AlO45- tetrahedra break down into smaller units, reducing the degree of polymerization, and the radius of ions decreases, thus

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where Vm is the molar volume of the slag and Vi* is the molar volume of the pure componet i, xi is the mole fraction of component i, and VE is the excess molar volume of the slag.

The molar volume of slags can be expressed by the following equation: [2] Here, xi is the mole fraction of component i, Mi is the molecular weight of component i, and ρ is the measured density of the slag.

We have to obtain the relation between the excess molar volume and composition. The regular solution approximation rule is most widely used (Shu, 2007). The excess molar volume can be expressed as follows: The Journal of The Southern African Institute of Mining and Metallurgy

Measuring and modelling of density for selected CaO-Al2O3-MgO slags [3] where aij represents the parameters for components i and j, which can be obtained by optimizing an appropriate amount of experimental data in a certain compositional and temperature range. xi and xj indicate the mole fractions of component i and j, respectively.

The average error of all calculated values can be assessed by using Equations [4] and [5]. δn is the percentage difference between the calculated and measured values. Δ(%)is calculated by taking the summation of all absolute values of δn and dividing by the total number of data.

Table II

Molar volumes of the pure components Oxide

Temperature (K) dependence of molar volume (m3/mol)

CaO Al2O3 MgO

20.7×(1+1×10-4×(T-1773))×10-6 (28.31+32xAl2O3-31.45x2Al2O3)×(1+1×10-4×(T-1773))×10-6 16.1×(1+1×10-4×(T-1773))×10-6

Table III

Values of model parameters for selected slags a12*

[4]

[5] The molar volumes of the pure oxides (Table II) recommended by Mills et al. were used in the present model.

Now that all necessary data has been collected, using the experimental data in the CaO-MgO-Al2O3 slag system and its subsystem, the optimized parameters are shown in Table III. The estimated values using the present model were compared with the experimental data and the data for the CaO-Al2O3 system (Dou et al., 2009; Slag Atlas, 1981; Ogino and Hara, 1977), MgO-Al2O3 system, and the CaO-MgOAl2O3 system (Slag Atlas, 1995) from the literature to verify the model. The comparison of the experimental data with the model calculated molar volumes are shown in Figure 6. It can be seen that the estimated values agree with the experimental

a13

a23

a123

CaO-Al2O3 system 0.47 MgO-Al2O3 system -5.42 CaO-MgO-Al2O3 system 0.47 2.93 -5.31 -202.00 * 1, 2 and 3 representative of CaO, Al2O3 and MgO, respectively.

data, and for all calculated values the average error Δ(%) is 2.20%. From the data for the CaO-MgO-Al2O3 slag system and its subsystem, the application of molar volume model confirmed that the present expanded approximation rules are applicable to predict the molar volume of the melts discussed.

Conclusions Both of the MgO content and the mass ratio of CaO/Al2O3 had an influence on the density of the selected slag. With a mass ratio of CaO/Al2O3 of unity, the density at first decreased with increasing the MgO content, following by an increase.

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Figure 6 – Comparison between experimental data and predicated values

Measuring and modelling of density for selected CaO-Al2O3-MgO slags Changes in the mass ratio of CaO/Al2O3, at a constant MgO content of 5.5 wt.%, gave rise to a similar density trend as changes in the MgO content.

An attempt has been made to estimate the molar volume of the CaO-Al2O3-MgO slag investigated in the present work.

Application of the molar volume model confirmed that the expanded approximation rules are applicable for predicting the molar volume of the melts discussed.

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Acknowledgements The authors gratefully acknowledge the financial support of the National Natural Science Foundation of China (50874072, 51204115), the Natural Science Foundation of Jiangsu Province of China (BK20130308), and the China Postdoctoral Science Foundation (2014M561710).

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Investigation of factors influencing blending efficiency on circular stockpiles through modelling and simulation

by Z. Loubser* and J de Korte†

The aim of a blending stockpile is to minimize the natural variation in properties of a material. Two simulation models were written, for coneshell and chevcon stacking, and used to investigate the influence of different stockpile parameters on blending efficiency. The coneshell model is used to show that blending efficiency is highly sensitive to stacked height, with lower output variances observed for higher capacities. Output variance in both models also decreased when the stacker movement increment was decreased. For chevcon stacking, increasing stacker speed significantly increases the blending efficiencies. Increasing the length of the blending tail also reduces output variation, but this parameter would need to be weighed against the accompanying decrease in buffer capacity.

Keywords stockpiles, blending efficiency, modelling, simulation, stacking mode.

Introduction Even in the most uniform deposit of materials in nature, material properties will vary. The degree of variation can be measured, and usually resembles a normal distribution. The aim of blending and homogenization operations is to minimize the standard deviation of this normal distribution (De Wet, 1994). In general, the reduction in variation will result in a more efficient and cost-effective downstream process. Cement plants need a precise blend of raw materials to ensure desired kiln performance; boilers used in power generation are optimized for fuels of a constant specification; and even though the separation density in a coal washing plant can be adjusted, a constant product quality cannot be produced from highly variable feed (Denny and Harper, 1962). The process efficiency, product quality, and environmental compliance depend on consistency of characteristics in the material fed (Kumral, 2005).

Many industries involved in minerals processing require a level of raw material homogenization as part of their process.

Product homogenization can also be needed if the process yields a product that is not sufficiently uniform (Van der Mooren, 1962).

One of the most widely used methods of homogenization is stockpiling.

Homogenization is not the only function of a stockpiling system. Stockpiles act as buffers The Journal of The Southern African Institute of Mining and Metallurgy

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Homogeneous blending on circular stockpiles Circular blending piles have been used worldwide since the late 1970s, notably in the coal, steel, and cement industries (Gerstel, 1989). On a circular stockpile (Figure 1) the stacker and reclaimer rotate about a central axis in the same direction, with one end being stacked while the other is reclaimed. The operation is therefore continuous, which is why circular stockpiles pose the advantage of not exhibiting the end-cone effect. This effect is observed in linear piles as a high degree of variation in material properties at the ends of the piles caused by an over-representation of the last-stacked material (Robinson, 2004).

* Partners in Performance International (previous, 2013 - BHP Billiton Energy Coal South Africa). † CSIR, South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Jan. 2015 and revised paper received Apr. 2015.

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Synopsis

between two processes, so that either can continue without being constrained by the other. More often than not, the buffering function is the primary economic consideration in stockpile design, so when homogenization options are evaluated only minor adjustments to the existing system can be considered (Robinson, 2004). If a consistently blended feedstock is provided, the following advantages can be expected (De Wet, 1983): ➤ Stable process operation, resulting in lower operating costs and higher product quality ➤ The process plant can be optimally sized for a given throughput rate, as the grade of feedstock can be more accurately predicted ➤ Product yield is higher, therefore raw material consumption (and consequently costs) is lower ➤ Product quality is controlled within smaller tolerances. This has cumulative advantages, as the product from one process is often used as input to another.

Investigation of factors influencing blending efficiency on circular stockpiles Circular stockpiles are stacked in one of two ways, either in coneshell or chevcon mode. Coneshell stacking is the most common method for storage systems that do not have a homogenization purpose (Wintz, 2011). The stacker starts discharging material at a single point, forming a cone. It then moves incrementally to form successive cones in shells over the first cone. Bond, Coursaux, and Worthington (2000) note that this method should not be used when a high degree of variability is present in the raw material feed.

In chevcon stacking, the stacker travels at almost constant speed back and forth within a preset angle of rotation, continuously discharging material (Figure 2). At the same time, the stacker moves up and down from the top of the existing stockpile to ground level, forming a blending tail.

The angle of incline chosen for this motion needs to be the same as, or smaller than, the angle of repose of the material, and the length of the blending tail is determined by homogenization requirements (Hurwitz and Ackermann, 1999).

Material is reclaimed by a full-face harrow reclaimer, scraping the cross-section of the pile at an angle approximately equal to the angle of repose of the material. Since the cross-section of the pile will consist of increments from different stacked layers, the discharged material is a blend of the stacked material. Reclaiming using a full-face reclaimer results in the highest homogenization efficiency (SACPS, 2011).

A homogeneously blended stockpile has the same composition throughout, and the properties of any smaller sample of material will apply to the pile as a whole.

Homogenization implies that the fluctuations of a property in the input flow are smoothed in the output, which results in a reduced standard deviation (Cerstel, 1989). The degree of homogeneity can be expressed as the standard deviation of a material property relative to its mean value (Denny and Harper, 1962).

According to Kumral (2006), the efficiency of a blending system is dependent on three factors: ➤ Stockpiling method. This is the way in which material is placed onto the stockpile by the stacker. The stacker is mostly responsible for homogenization, not the reclaimer (SACPS, 2011) ➤ Stockpiling parameters. These include the length, width, number of layers stacked, equipment properties of the stacker and reclaimer, and raw material characteristics, among others ➤ Input variability. The frequency and amplitude of the variation of material properties in the input stream.

A stockpile is successful in homogeneous blending if the instantaneous analysis of reclaimed material closely resembles the average value of the whole pile. This, according to De Wet (1994), will depend on the storage capacity of the bed, the nature of variation in the input material, and the degree of quality control exercised. According to Bond, Coursaux, and Worthington (2000), blending efficiency can be increased by increasing the volume of the stockpile or increasing the stacking speed (i.e. stacking more layers).

Determining blending efficiency One way of measuring blending or homogenization efficiency is the variance reduction ratio (VRR), where: [1] In Equation [1], σ2out and σ2in are the output and input variances respectively. Input and output variances should be calculated on the same base, i.e. identical weights or volumes (Kumral, 2006).

De Wet (1994) stated that, subject to certain statistical assumptions, the homogenizing effect of a blending pile can be estimated as: [2]

Figure 1 – Model representation of a circular pile being stacked and reclaimed in coneshell mode

where S is standard deviation and N is the number of layers the reclaimer cuts into. In the second part of the equation, V is the stacker travel speed in m/min, F is the stockpile crosssectional area in m2, and Q is the stacking rate in m3/h.

Previous work on stockpile modelling and simulation

Figure 2 – Model representation of a circular pile being stacked in chevcon mode

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Although much work has been done on the simulation of blending (Robinson, 2004; Gerstel, 1989; Hurwitz and Ackermann, 1999; Kumral, 2006; Gerstel, 1979; Pavloudakis and Agioutantis, 2003), it has been focused mostly on linear stockpiles. Modelling is often of a statistical nature, which requires some prior knowledge of the input grade variation.

Statistical models are also based on many assumptions, some of which are not applicable to most material-handling applications.

Gerstel (1989) developed a mathematical model, based on statistical approximation, to offer a reasonable estimation of The Journal of The Southern African Institute of Mining and Metallurgy

Investigation of factors influencing blending efficiency on circular stockpiles the homogenization that can be obtained by using a circular blending pile if constant stacker feed is assumed. Hurwitz and Ackermann (1999) registered a patent for a bedblending optimization model that created a database of the grades of materials in the pile and predicted the aggregate output composition of vertical sectors when they are reclaimed. Important stockpile modelling work was done by Robinson (2004), who used statistical and geometrical modelling to predict the expected variation in output properties from different linear stockpile configurations. The model used a matrix of proportions representing the amount of material from a layer that is contained in an inclined reclaimer slice, again assuming constant stacker feed. Kumral (2006) used Robinson’s modelling principle to build a stockpile simulator, modelling a stockpile as a set of blocks created by the intersection of reclaimed slices and stacked layers. The simulator is used to build a multiple regression model for optimizing stockpile parameters by minimizing VRR for different scenarios. Pavloudakis and Agioutantis (2003) developed a simulation model that uses properties of stacked material combined with stockpile geometry to predict the output from a linear stockpile. The end-cone sections were ignored, and only the triangular prism-shaped part of the pile was modelled.

No published simulation model could be found that accurately estimates the blending efficiency and output characteristics of a circular stockpile when little is known about the input characteristics. The available literature also lacks a comparison between the blending efficiencies that can be achieved for chevcon and coneshell stacking methods, although this comparison has been investigated repeatedly for linear stacking methods.

Simulation model development From a perusal of the literature available it is clear that the simulations documented so far used a degree of assumption and approximation in the model development. This will, of

course, always be the case in modelling and simulation.

Although these approximations are, for the most part, an accurate enough indication, they could be adjusted to better reflect reality. The development of this simulation model aimed to eliminate two important assumptions. Firstly, almost all of the published simulations that were found use a vertical ‘slice’ to model the action of a full-face reclaimer. In reality the reclaimer is cutting into the stockpile at an angle, therefore the proportions of each stacked layer represented in a reclaimed section will be significantly different from those predicted for a vertical cut. Secondly, most of the simulations assume a constant material flow from the stacker, which is rarely the case in practice. Modelling of the stacked layers can be made more accurate by accommodating variation in the material feed rate.

Stockpile modelling, as discussed in this framework, can be used to predict any additive material property, but in most cases consistent ‘grade’ of material will be the main driver for plant optimization. ‘Grade’ can refer to any one of a range of material properties, depending on the material being processed and the downstream application.

The principle behind the model design is that the entire stockpile volume can be envisioned as being made up of many small blocks. This echoes the ideas of Robinson (2004) and Kumral (2006), but is approached in an entirely different way. Instead of stacking individual blocks of set volume, and reclaiming them in another order to achieve a blend, a Cartesian coordinate system is used to model the placement and properties of each block.

A three-dimensional grid can be drawn; large enough for the entire stockpile to fit inside. The grid corresponds to three Cartesian axes, with the x- and y-axes forming the base of the stockpile and material being stacked upwards in the increasing z-direction. Therefore every block in the grid has an address, which can be represented by using the position of the block in the Cartesian coordinate system.

The grid is divided so that blocks have side lengths of 1 m,

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Figure 3 – Two-dimensional representation of how blocks are filled

Investigation of factors influencing blending efficiency on circular stockpiles

Figure 4 – User interface for coneshell model

thus each block has a volume of 1 m3. This is equal to roughly one ton of material, depending on the material type and bulk density. Since feed rates to stockpiles are in the order of 1000 t/h, it is reasonable to assume minimal variation of grade within each of these blocks. When material is stacked, the blocks fill up according to stockpile geometry, with each block being assigned the grade of the stacked increment. When blocks are later reclaimed, the grades of the blocks in each reclaimed ‘slice’ are aggregated to generate predictions of grade variation in the stockpile output.

Visual Basic for Applications (VBA) was used to develop two simulation models, one each for circular stockpiles stacked in coneshell and chevcon mode. Sheets are provided for data entry, and a central user interface (Figure 4) controls the simulation parameters.

Model validation In order to prove the model’s validity, its predictions are compared to known outputs. For this purpose data was collected from a South African coal processing facility. The simulation model was used to predict the output grade variation of a stockpile, which was compared to the actual grades recorded. The case study facility uses a circular product stockpile with a live capacity of 45 000 t, and material is sampled before it is stacked and after it is reclaimed, at 2-hour intervals. The data-set included information on tonnage per hour and grade results from samples.

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Two different data-sets were used for simulations, each generating hour-on-hour predictions of the stockpile running continuously for 1 month, using stacker feed rates and stacked grades as input. The predictions were then compared to real production data from the given month, recorded by the reclaimer sampler. The validation process indicated that the simulation models provide reasonably accurate predictions of grade variation in the stockpile output. Despite challenges to validation, such as missing data and output contamination, the predicted values fell within 5% of the recorded grades.

The models were deemed suitable to be used for comparison of stockpile parameters in the interest of increasing blending efficiency.

A detailed description of the model development and validation is provided in the unpublished thesis preceding this paper (Loubser, 2012).

Simulations performed Once it is confirmed that the simulation model offers a reasonable approximation of reality, the model can be used to evaluate the influence of some operational parameters on the blending performance of the stockpile. This is a useful function of a simulation model; the effect of changes can be investigated without actual implementation, and potential negative outcomes can be avoided. Users can therefore optimize their process through risk-free experimentation.

Table I details properties of the stockpile used in the simulations to follow.

The Journal of The Southern African Institute of Mining and Metallurgy

Investigation of factors influencing blending efficiency on circular stockpiles Stockpile properties used for simulation Property type

Property

Variable name

Value

Unit

Material

Repose angle Bulk density Diameter Stacker radius Max. height

aRepose bulkDens diam rStack hStockpile

38 0.9 110 28 19

degrees t / m3 m m m

Stockpile

Table II

Coneshell simulations Coneshell stacking Stacker increment (degrees)

Stacked height (m)

10 6 8 12 14 10 10 10 10

17 17 17 17 17 15 16 18 19

Simulation 1 (base case) Simulation 2 Simulation 3 Simulation 4 Simulation 5 Simulation 6 Simulation 7 Simulation 8 Simulation 9

Table III

Chevcon simulations Chevcon stacking

Simulation 1 (base case) Simulation 2 Simulation 3 Simulation 4 Simulation 5 Simulation 6 Simulation 7 Simulation 8 Simulation 9 Simulation 10 Simulation 11 Simulation 12 Simulation 13

Angle of rotation (degrees)

Stacker speed (m/min)

Layers stacked

100 50 75 125 150 100 100 100 100 100 100 100 100

25 25 25 25 25 20 22.5 27.5 30 25 25 25 25

200 200 200 200 200 200 200 200 200 100 150 250 300

The starting condition for each simulation was an empty stockpile with no material present.

For the coneshell model two variables were chosen for the purpose of investigation. The increment that the stacker moves between successive cones can be varied and could possibly influence the efficiency of blending. The variation of blending efficiency with changes in stacked height was also investigated. Table II shows the list of coneshell simulations performed.

The influences of three variables were investigated for the chevcon model. These were the length of the blending tail The Journal of The Southern African Institute of Mining and Metallurgy

(i.e. the angle of stacker rotation), the stacker movement speed, and the number of incremental layers contacted by the reclaimer when cutting into the stockpile cross-section. When the number of layers is increased, the increment that the stacker moves between consecutive layers is decreased.

Therefore varying this parameter can also be seen as varying the stacker step increment. Details of the experimental simulations performed on the chevcon model are shown in Table III. Once all simulations are complete, the results are compared on the basis of blending efficiency (VRR) in order to assess the influence of each variable. The results obtained for the chevcon and coneshell methods are also compared in order to quantify the difference in potential blending efficiency achieved by using either method.

Simulation results Findings from the simulation results are discussed in this section, and further related to work by other authors in the conclusions.

Coneshell model As noted in the previous section, two parameters were evaluated in the coneshell stacking model.

Figure 5 shows how the VRR varies in relation to the stacker movement increment. As the distance the stacker moves between cones is increased, the VRR can be seen to increase almost linearly. This is an expected result, as smaller increments between cones is analogous to more layers being stacked in any given space, which would result in better blending efficiency. The increment between cones should thus be minimized as far as possible. Another important result from Figure 5 is that as the stacker increment tends to zero, the VRR tends to 0.25, implying a minimum achievable VRR. Figure 6 shows how VRR varies with stacked height.

Although the data points resemble a straight line, a linear relationship is not possible. If VRR were to decrease linearly with an increase in stacked height, at a point before 25 m stacked height the line would cross the x-axis, implying zero

Figure 5 – Coneshell model – VRR vs stacked increment VOLUME 115

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Table I

Investigation of factors influencing blending efficiency on circular stockpiles variance. Since completely consistent output from a blending pile is not practically achievable, it makes more sense to approximate the relationship between VRR and stacked height as an exponential function.

An increase in maximum stacked height of an existing pile will rarely be possible, since the diameter of the stockpile would have to increase accordingly. Insights about the influence of pile capacity on blending efficiency can, however, be useful when weighing the advantages of different stockpile designs.

Variation in output grades as generated by the coneshell model has proven to be highly dependent on stacked height, with only a slight dependence on changes in the stacker movement increment.

Chevcon stacking The influence of three different variables on the expected output variance from the chevcon model was investigated.

The results are displayed and discussed below.

Gerstel (1989) stated that in order for sufficient blending to take place on a chevcon-stacked circular blending pile, the tail length had to be at least four times the horizontal length of the reclaimer harrow, i.e. the horizontal distance between points A and B in Figure 7.

Figure 8 demonstrates how output variance decreases

with increasing blending tail length. The relationship between VRR and tail length can also be approximated as an exponential function. According to Gerstel’s rule of thumb, the blending tail for the stockpile in question would need to be at least 87 m, but Figure 8 would suggest that increasing the blending tail beyond 60 m would not yield a significant further decrease in variance. When evaluating this parameter one must also keep in mind that increasing the size of the blending section decreases the stockpile storage capacity.

Incremental improvements in grade consistency must thus be weighed against buffering needs.

From the work of Bond, Coursaux, and Worthington, (2000), an increase in stacker speed is expected to yield a decrease in output variance, as faster stacker movement means that material from each time increment is stacked into more layers. Figure 9 shows that when VRR is plotted as a function of stacker speed, the data can be fitted to a parabolic

Figure 8 – Chevcon model – VRR vs blending tail length

Figure 6 – Coneshell model – VRR vs stacked height

Figure 7 – Reclaimer harrow

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Figure 9 – Chevcon model – VRR vs stacker speed VOLUME 115

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Investigation of factors influencing blending efficiency on circular stockpiles Table IV

Results of coneshell vs chevcon comparison Variance

VRR

μ − 2σ

μ + 2σ

22.71 22.71

0.202 0.075

0.232 0.121

21.81 22.16

23.61 23.26

curve. Although this would imply that values of stacker speed exist for which variance equals zero, which has been established as a practical impossibility, the parabolic shape seems a viable result for this section of the graph. The curve suggests that for low stacker speeds there is little difference in the resultant variance, but there exists a critical speed after which variance decreases dramatically. From this result it can be concluded that, as far as is practically viable, stacker speed should be maximized if a consistent output grade is desired.

The number of chevcon layers that contact the reclaimer face was related to the stacker step increment as shown in Equation [3].

[3] The number of layers was used to specify the simulation parameters, but was exchanged for step increment in discussion of the results, as this is a more relatable parameter.

Figure 10 shows how VRR changes with increasing stacker movement increment between successive chevcon layers being stacked. The graph does not seem to follow a recognizable form, but some conclusions can be drawn. It appears that for very low values of the stacker increment (smaller than one degree) the model delivers inaccurate results. This could be because the layer increment is not sufficiently large in comparison to the calculation increments used by the modelled stacker and reclaimer, which would indicate that some of the approximations used are insufficient. These inaccuracies induce false variance in the results. A solution to the problem would be to use smaller calculation increments to model the behaviour of the

Figure 10 – Chevcon model – VRR vs stacker increment The Journal of The Southern African Institute of Mining and Metallurgy

stockpile for very small values of stacker increment. For values of stacker increment greater than or equal to one degree, the relationship to VRR has a more expected form: variance is reduced if smaller increments of stacker movement are used (as concluded for the coneshell model).

For this situation it appears that the optimal value of stacker increment is one degree, but the result would need to be investigated further by running the model on a PC with more processing power.

A reduction in variation of output grades from the simulated chevcon stockpile has proven to be highly dependent on the stacker speed. Although increasing the blending tail length would increase blending efficiency, significant changes would be seen only in the lower ranges of tail length, i.e. there exists a limit after which negligible additional benefit will be seen from an increase in tail length.

Unfortunately, no conclusions could be made about the level of dependence of the chevcon model on stacker movement increment.

Coneshell vs chevcon stacking In order to compare the output variance generated by the coneshell and chevcon methods for the same input, one additional simulation was run for each model. These simulations used the optimal values of each of the parameters as identified in the first part of this section. For the coneshell model, a stacker movement increment of six degrees was combined with a stacked height of 19 m. In the chevcon simulation, the stacker rotated about an angle of 150 degrees for every layer, stacking speed was set to 30 m/min, and a stacker increment of one degree was used. The results are shown in Table IV. For the coneshell model a VRR of 0.232 is obtained. This is not significantly lower than most of the other coneshell stacking results, which seems to imply that variation of the stockpile parameters has a limited influence on the blending efficiency for a stacker in coneshell mode. The result leads to the conclusion that, for the stockpile used in simulations, VRRs of less than 0.2 are unlikely to be achieved while stacking in coneshell mode. For the chevcon stacking mode, however, a large reduction in variance is obtained by optimizing the stacker parameters. The very low output variance indicates that, for the right stockpile configuration, near-perfect blending can be achieved by stacking in the chevcon mode. The result will, of course, be subject to induced variances, like that caused by sampling (Robinson, 2004).

From the results discussed, chevcon stacking has proven to deliver much better consistency in the output grades reclaimed from a circular stockpile. As reported by Bond, Coursaux, and Worthington, (2000) and Wintz (2011), coneshell stacking is best reserved for when homogenization is not an important requirement.

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Coneshell stacking Chevcon stacking

Mean

Investigation of factors influencing blending efficiency on circular stockpiles Conclusions Simulation models have been developed for coneshell and chevcon stacking, and used to investigate the sensitivity of blending efficiency on circular stockpiles to various input parameters. It was found that for coneshell stacking, a smaller increment between stacked cones will result in lower output variance, with a close to linear relationship being observed. The results also suggested that a minimum output variance exists, which corresponds to a VRR of approximately 0.25 for the range of parameters investigated. The influence of stacked height on blending efficiency was also evaluated for the coneshell stacking model. It was found that in order to minimize output variance the stacked height must be maximized. The impact of this operating variable will be limited, as the stacking height of an existing stockpile cannot be arbitrarily increased. In order to increase stockpile height, a new stockpile must be designed. Both of these results confirm findings from the literature. Denny and Harper (1962), De Wet (1983, 1994), and Petersen (2004) all noted that increased blending efficiency can be obtained by stacking more layers or increasing the stockpile capacity.

Stacking more layers in coneshell mode is analogous to decreasing the distance between cones, while an increase in stacked height would result in an increase in stockpile capacity.

For a chevcon-stacked pile a decrease in output variance is observed for longer blending tails. The results suggest, however, that beyond a blending tail angle of about 125 degrees no significant decrease in VRR will be obtained. It is important to weigh the increased blending efficiency against the decrease in available stockpile volume when considering this factor. A rotation angle of 125 degrees means that more than a third of the stockpile volume will not be available as buffer capacity. Simulation results for varying stacker speeds indicate that there exists a critical speed of about 25 m/min for the simulated stockpile, above which a large reduction in output variance can be achieved. Therefore the stacker speed should be maximized as far as is practical. The last parameter that was evaluated is the increment that the stacker moves between chevcon layers. For very small values of the movement increment, smaller than 1 degree, the model proves to be inaccurate. This could be because the layer increment is not sufficiently large in relation to the calculation increments used by the stacker and reclaimer, which means that some of the approximations used would prove to be insufficient. For movement increments greater than 1 degree, an increase in layer increment results in an increase in output variance.

In order to compare the output variance generated by the coneshell and chevcon stacking methods for the same input, additional simulations were run for each model. The simulations used optimal values of stockpile parameters as identified in the sensitivity analysis. It was found that for coneshell stacking, optimization of the input parameters yields little improvement in blending efficiency, as the reduction in output variation that can be obtained is limited.

It was concluded that, for the stockpile simulated, VRRs of less than 0.2 would unlikely be achieved while stacking in coneshell mode. This is an expected result, as Bond, Coursaux, and Worthington, (2000), Erasmus (2001), and Wintz (2011) all report high output variation for the

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coneshell stacking mode, and deem it unsuitable for applications requiring high blending efficiency. The output variance generated by the chevcon model was half of that achieved with coneshell stacking, a VRR of less than 0.1, proving that the investment in chevcon stacking infrastructure is one that every production facility seeking improvement in output grade consistency should consider.

This also confirms Bond’s approximation of a 10:1 reduction in variance for a stockpile stacked in chevcon mode and reclaimed with a full-face reclaimer (Bond, Coursaux, and Worthington, 2000).

References BOND, J.E., COURSAUX, R., and WORTHINGTON, R.L. 2000. Blending systems and control technologies for cement raw materials. Industry Applications Magazine, vol. 6, no. 6. pp. 49–59.

DE WET, N. 1983. Homogenizing/blending plant applications in South Africa with special reference to Gencor's Hlobane and Optimum plants. Bulk Solids Handling, vol. 3, no. 1. pp. 49–59.

DE WET, N. 1994. Homogenizing/blending in South Africa - an update. Bulk Solids Handling, vol. 14, no. 1. p. 93.

DENNY, R.J. and HARPER, W.G. 1962. The homogenisation of raw coal. Fourth International Coal Preparation Congress, Harrogate, UK, 28 May–1 June 1962.

ERASMUS, J.H. 2001. Bulk raw materials storage selection. Beltcon. South African Institute of Materials Handling.

GERSTEL, A.W. 1979. The homogenisation of bulk material in blending piles.

Unpublished thesis, Delft University of Technology, The Netherlands.

GERSTEL, A.W. 1989. Blending in continuous piles. Bulk Solids Handling, vol. 9, no. 4. p. 413.

HURWITZ, M.J. and ACKERMANN, R.E. 1999. Real-Time Optimization for Mix Beds.

European patent EPO943565 A3. Gamma-Metrics KUMRAL, M. 2005. Quadratic programming for the multivariable pre-homogenization and blending problem. Journal of the South African Institute of Mining and Metallurgy, vol. 105, no. 5. pp. 317–322.

KUMRAL, M. 2006. Bed blending design incorporating multiple regression modelling and genetic algorithms. Journal of the South African Institute of Mining and Metallurgy, vol. 106, no. 3. pp. 229–236.

LOUBSER, Z. 2012. A decision support tool for circular stockpile management using simulation. Unpublished thesis. Stellenbosch University, Stellenbosch, South Africa.

PAVLOUDAKIS, F.F. and AGIOUTANTIS, Z. 2003. Simulation of bulk solids blending in longitudinal stockpiles. International Journal of Surface Mining, Reclamation and Environment, vol. 17, no. 2. pp. 98–112.

PETERSEN, I.F. 2004. Blending in circular and longitudinal mixing piles.

Chemometrics and Intelligent Laboratory Systems, vol. 74, no. 1. pp. 135–141.

ROBINSON, G.K. 2004. How much would a blending stockpile reduce variation?

Chemometrics and Intelligent Laboratory Systems, vol. 74, no. 1. pp. 121–133.

SACPS. 2011. Materials handling. Coal preparation in South Africa. England, T., Hand, P.E., Michael, D.C., Falcon, L.M., and Yell, A.D.(eds.). South African Coal Processing Society, Pietermaritzburg..

VAN DER MOOREN, A.L. 1962. The homogenisation of streams of coal. Fourth International Coal Preparation Congress, Harrogate, UK, 28 May–1 June 1962.

WINTZ, S. 2011. Bed storage and blending applied to gypsum. Global Gypsum. http:// http://www.globalgypsum.com/magazine/articles/432-bed-storageand-blending-applied-to-gypsum. ◆ The Journal of The Southern African Institute of Mining and Metallurgy

http://dx.doi.org/10.17159/2411-9717/2015/v115n8a16

Investigation into strata behaviour and fractured zone height in a high-seam longwall coal mine

by G. Song* and S. Yang*

Highly mechanized mining equipment for longwall faces was first introduced to China in 1978. However, preliminary trials were not successful due to equipment capacity limits and roof control problems. The next few decades, however, witnessed great progress in coal production, mining heights, and panel widths, which were due to improvements in mining techniques and equipment, especially the hydraulic-powered shield (Wang, 2013).

The maximum annual production in a single panel stood at less than 2.5 Mt in 1998, but this increased dramatically to 12 Mt in 2008 and 14 Mt in 2011 (Yuan et al., 2010). In 2007, Shangwan coal mine developed the world’s first 6.3 m single-pass high-seam longwall face. The mining height increased to 7 m in Bulianta mine in 2009. The trend of increasing shield maximum support height from 2000–2014 is shown in Figure 1, and Figure 2 shows the shield used in Wangzhuang mine. Table I lists the parameters of the panels and equipment, which compares the improvement of these faces from 6.3 m to 7 m. The increase of mining height has led to nearly 10% improvement in recovery and around 1.2 Mt of additional annual production, but requires heavier and larger installed-power equipment.

Synopsis The development of techniques and equipment for single-pass high-seam longwall mining in China is reviewed. Some methods used to obtain the fractured zone height are discussed. The boundary between the caved and fractured zones is generally not clear, but one important difference that distinguishes the two zones is that horizontal compressional forces exist only between blocks of fractured zone. Based on this, a theoretical method is presented to investigate the destabilizing modes of the main roof (by sliding or rotation) and is used to determine if the main roof is in the caved or fractured zone. This method considers the mining height, immediate roof thickness, bulking factor, main roof thickness, main roof strength, main roof periodic weighting interval, and vertical stress in the overlying strata. To verify the method, a representative physical model of Wangzhuang coal mine is developed. The movement of overlying strata and fractured zone height are thus obtained. According to the results, the total collapsed height of the strata reaches about 70 m above the coal floor; the first main roof bending interval is 45 m and periodic bending distance is 10–15 m; the maximum strata subsidence is around 62 mm; and the presence of a three-hinged arch causes the fluctuation of subsidence in the same stratum level. Based on the theoretical and experimental analysis, the middle of the main roof is considered the boundary between caved and fractured zone; the caved and fractured zone heights are about 21 and 49 m respectively.

Keywords thick coal seam, single-pass longwall mining, physical modelling, bulking factor, caved zone, fractured zone, roof strata behaviour.

Physical modelling investigation

Thick coal seam mining in China Thick coal seams (>3.5 m) account for about 44% of the coal rescource/reserves and more than 40% of the total annual production in China (Zhao 2004; Wang 2009). Three mining methods – slice mining, top coal caving, and single-pass high-seam longwall mining – are frequently used in underground thick-seam operations. Compared with slice mining and top coal caving, high-seam single-pass longwall mining significantly simplifies the mining processes, reduces roadway excavation, and improves coal production, productivity, and recovery. The favourable economics, high production, and technical simplicity make high-seam single-pass longwall mining one of the most popular underground mining methods for thick coal seams in China.

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The fractured zone height and roof control are the most important issues in single-pass highseam longwall mining in China. It is believed that strata movement is likely to be more severe; face failure and roof falls more frequent (Yuan et al., 2010), and the caved/fractured zone height higher due to larger mining thickness and high-intensity mining activity (Sun et al., 2013). Generally, physical modelling is used to study the dynamic movement process in overlying strata during mining. This model can record the roof

* College of Resources and Safety Engineering, China University of Mining and Technology, Beijing, China. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received May 2015 and revised paper received June 2015.

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Introduction and background

Investigation into strata behaviour and fractured zone height in a high-seam longwall coal mine research can be used to verify numerical findings. Yang et al.

(2009) investigated the influence of mining thickness on abutment pressure and three zones (caved zone, fractured zone, and continuous subsidence zone over goaf) in a physical model experiment. The results showed that the area influencd by abutment pressure and the height of the caved/fractured zone increase as mining height increases.

The floor movement can also be observed in physical simulation experiments. Jiang et al. (2011) studied the failure of coal seam floors above confined water experimentally and analytically. The displacement of floors was recorded, indicating that the floors move upwards and the displacement increases as the face advances. Three zones in the floors (mining-induced fissure zone, water-resisting zone, and confined water ascending zone) were observed. Zhao et al.

(2013) developed this study further by applying horizontal loads to the experimental model according to the ground stress survey. They concluded that the largest upward movement occurs 30 m behind the panel, and the closer the floor is to the aquifer, the larger the displacement would be.

Panel 8101 at Wangzhuang is the first high single-pass longwall face in this mine, with no previous strata behaviour data available. Thus obtaining this data is of great significance for better control of the roof and for successful mining.

A theoretical analysis and a physical model to obtain the height of caved/fractured zone and Wangzhuang and describe roof movement behaviour are presented in this paper.

Figure 1 – Maximum height of shield from 2000

Theoretical analysis of caved, fractured zone, and strata behaviour Previous work Strata movement and ground subsidence have been investigated in detail by researchers around the world.

Wagner et al. (1991) reviewed the ground subsidence at South African collieries during total-seam underground extraction. The vertical subsidence in South African mines was found to be less than at European mines, due to the presence of competent layers in the overlying strata. Singh et al. (1999) monitored strata parameters and the behaviour of

Figure 2 – Type ZY15000/33/72D shield used at Wangzhuang

failure development, strata subsidence, and the structure of roof blocks. Li et al. (2005) compared the results of numerical and experimental studies of strata movement above the collapsed gob and suggested that experimental

Table I

Comparison of 6.3 m and 7.0 m panels Colliery

Shangwan

Bulianta

Bulianta

Shangwan

Panel built (year) Panel number Panel height, m Panel width, m Panel length, m Recovery Production, Mt Shield type Shield load density, MPa Shield total weight, t Shearer type Shearer mining range, m Shearer total installed power, kW Shearer total weight, t Conveyor type Conveyor total installed power, kW Conveyor carrying capability, t/h

2007 51202 6.3 301 4466 87.1% 10.87 ZY10800/28/63 1.05-1.12 45 SL1000 2.5-6.2 2×1080 135 48 mm AFC% 3×1000 —

2007 32301 6.3 300 5220 88.7% Around 11 ZY10800/28/63 1.05-1.12 45 SL1000 2.5-6.2 2×1080 135 48 mm AFC% 3×1000 —

2009 22303 7.0 301 4966 96.7% Around 12 ZY18000/32/70 1.39-1.43 69 SL1000 (7 m) 2.7-7 2590 110-155 PF6H-1542 3×1600 6000

2010 12205 7.0 319 4231 Around 97% Around 12 ZY18000/32/70 1.39-1.43 69 SL1000 (7 m) 2.7-7 2590 110-155 PF6H-1542 3×1600 6000

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The Journal of The Southern African Institute of Mining and Metallurgy

Investigation into strata behaviour and fractured zone height in a high-seam longwall coal mine powered support for exploitation of a 7.5 m sublevel caving face in India through field observation. The movement of roof strata was visualized in a simulated physical model. Cui et al.

(2000) used nonlinear geometric field theory to predict the surface subsidence induced by underground mining. Ju et al.

(2013) investigated the structural characteristics and behaviour of key/competent strata in the overburden of a 7 m high panel through in-situ observation and a laboratory physical model.

Generally, the collapsed strata above the goaf can be classified into three zones; the caved zone, fractured zone, and continuous subsidence zone (Qian and Shi, 2003). It is important to know the fractured zone height and boundaries of these zones, especially for mines with roof water inrush potential. Immense amounts of field monitoring work have been conducted in China, and empirical equations to calcalate fractured zone height have been derived, shown as Equations [1] and [2] (State Bureau of Coal Industry, 2000).

[1]

is main roof thickness, and Δ is the void between caved immediate roof rock mass and overhang main roof), then the main roof is included in the fractured zone. Obviously, the void between caved immediate roof and overhang main roof is closely related to main roof breakage. In the case of a smaller void (Figure 3b), rotation is more likely to occur due to the support of waste rock on the floor before main roof blocks detach completely. A larger void is more likely to lead to main roof sliding. The main roof is then classified in the caved zone, with no compressive force formed between roof blocks (Figure 3c).

Bulking factor (Kρ), the ratio of broken rock bulk to original rock bulk, is used to describe the property of the rock mass (see Equation [3]). Das (2000) believed that the bulking factor controls the ground subidence. He classified superincumbent strata over a longwall panel into caving zone, unfilled void, freely detached zone, weighting zone, and stable superincumbent roof zone, and demonstrated that 45–60% of the total caving height is filled with bulk volume, leaving the remaining 40–55% as unfilled void that causes subsidence through:

[2] Based on empirical equations, the heights of the caved and fractured zone at Wangzhuang mine are 10.8–15.2 m and 40.5–51.6 m, respectively.

Since then, many investigators have carefully reviewed and expanded the research on fractured zone height. Xu (2009, 2012) and Wang (2013) believed that the position of the key stratum is closely related to fractured zone height. Hu (2012) developed the factors influencing fractured zone height (incorporating mining height, strata, panel length, seam depth, and panel advancing speed) and obtained a formula using multiple regression analysis to calculate fractured zone height. Kang (2009), Wu (2012), and Sun (2013) measured the fractured zone height using different methods such as borehole camera, drilling fluid, and drilling resistivity.

Theoretical solution

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Figure 3 – Main roof breakage styles VOLUME 115

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To obtain a specific height of the caved and fractured zones, it is necessary to clearly identify the boundary of these two zones. In general, the immediate roof is essentially soft and weak mudstone, which caves in after the shields are advanced; whereas the main roof – the massive, thick, and strong roof above the immediate roof – will bend, sag, split, rotate, or detach after a considerable distance of overhang for a long period (Das, 1999; Qian, 2003). The breakage modes of the main roof include rotation and sliding. The broken main roof blocks could be assigned to the caved or the fractured zone, depending on the horizontal compressional force between broken roof blocks. If rotation occurs in the main roof, it is likely that a horizonal compressional force is generated between roof blocks and this main roof is thus incorporated in the fractured zone. On the other hand, the main roof is included in the caved zone if sliding is seen, implying that no compressional force was exerted. Figure 3(a-c) shows the rotation and sliding of the main roof during first and periodic weighting. Figure 3d shows the threehinged arch formed in a rotation of the main roof.

The breakage style of the main roof thus determines whether the main roof is assigned to the caved or the fractured zone. Hou (2003) believed that if h>1.5·Δ(where h

Investigation into strata behaviour and fractured zone height in a high-seam longwall coal mine [3] where v0 is the original bulk of rock, and v1 is the bulk of the collapsed rock mass.

During mining, the weak immediate roof caves in and the broken rock is randomly deposited on the floor. Equation [4] applies if the collapsed immediate roof rock mass completely fills the gob. In this case, no void is left.

[4] where M and ∑h are the mining height and roof thickness before collapse, respectively.

Otherwise, unfilled void (Δ) will be left between the collapsed rock mass and main roof:

Experimental solutions

[5]

Physical modelling is often used to visualize the behaviour of the overlying strata by simulating the field conditions in a model constructed from similar materials. In general, the most important similarities between the model mine and the actual mine should be adhered to if reliable results are to be obtained (Jiang et al., 2011; Zhao et al., 2013).

[6]

Geometric similarity:

Equation [6] describes the sinkage of main roof.

The magnitude of unfilled void (Δ) and main roof sinkage (s) can be compared to determine the style of main roof breakage. If s > Δ (Equations [7] and [8]), roof blocks are more likely to rotate, and blocks are supported by the collapsed rock mass before they fully detach. s>Δ

angle of 3°–7°. There are five interbedded dirt bands in the coal seam, with a total thickness of 0.78 m; the thickest band reaches 0.4 m thickness in part of the seam. The width of the longwall face is 270 m, and the length is 546 m.

Figure 4 depicts the general geological section of Wangzhuang coal mine. The 14.4 m immediate roof above the coal seam is weak, soft mudstone, and the 8.8 m main roof is fine-grained sandstone. The 5 m immediate floor and 2 m main floor are mudstone and fine-grained sandstone, respectively. The Protodikonov’s Hardness Coefficient of the coal is 1.2. The coal type is excellent power coal, with low ash, sulphur, and phosphorus contents and a high calorific value.

The simulated equivalent material mine model was built on a 1:100 geomechanical scale. The ratio of model strata thickness to mine-scale strata thickness in the field is AL=1/100.

[7]

where AL is the geometric similarity ratio, and Lm and Lp are the thickness of model strata and full-scale strata, respectively

[8]

Time similarity:

Equivalently:

Accordingly, the roof is more likely to rotate and be classified in the fractured zone if the immediate roof thickness (∑h) is larger; or mining height (M) is smaller; or the main roof first weighting/periodic distance (L) is larger; or the bulking factor (Kρ) is larger.

Equations [9] and [10] show the main roof first and periodic weighting distances.

where AT is the time similarity ratio, Tm and Tp are the mining time in the model and in field mining activity. AT is 1/10 in this case.

[9]

[10] where L and L0 are the first and periodic roof weighting distance, respectively; RT is main roof tensile strength; q is the vertical stress in the overlying strata (Qian and Shi, 2003). In other words, a thicker or stronger main roof or a shallower seam would lead to a larger first caving distance and hence involve the main roof in the fractured zone.

Physical modelling Geomining conditions of Panel 8101 The representive physical simulation is based on Panel 8101 at Wangzhuang, which is worked by single-pass longwalling for the total extraction of a 6.3 m thick coal seam at 350–450 m depth. Panel 8101 inclines toward the southwest at an

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Figure 4 – General geological section The Journal of The Southern African Institute of Mining and Metallurgy

Investigation into strata behaviour and fractured zone height in a high-seam longwall coal mine Unit weight similarity:

where Aγ is the unit weight similarity ratio, and γm and γp are the unit weight of the model strata and field strata. Here Aγ is 1/1.7

Elasticity modulus and strength similarity:

where AE and Aσ (1/170) are elasticity modulus and strength similarity.

In this physical model, similar materials consist of aggregate (fine sand) and binders (gypsum, lime, and clay).

The final physicomechanical properties depend on the proportions of these materials, hence, duplicate experiments were repeatedly conducted to obtain the correct proportions.

Table II lists the material physicomechanical parameters of strata and the model dimension parameters.

The overall dimensions of the built model are 1 m × 1.2 m × 0.2 m. The model in Figure 5 shows the layout of the measuring points (small pins). Movements of these points were measured by an electronic theodolite and a digital camera before extraction, and were repeatly measured during every stage of extraction. The coal in the model is mined every 50 mm, which represents 5 m in mine scale.

immediate roof, 2 m in thickness (see Figure 6b). The first major fall occurs with a further 10 m advance (Figure 6c). The collapsed height in the overlying strata goes up to 21 m (starting from the bottom of the coal seam). These falling rocks are unable to fill the gob because of the large mining height.

(2) Main roof first bending At 45 m of face advance, the limiting span of overhang of the main roof is met and the first roof weighting initiates (Figure 6d). Both rotation and sliding occur. The collapsed height reaches 32 m. A three-hinged arch is formed between the broken main roof blocks. The collapsed main roof blocks compact the falling immediate roof on the floor.

(3) Main roof periodic bending After another 10 m of overhang, the main roof periodic weighting occurrs. Advance of the face to 55 m and 65 m generates periodic weighting (Figure 6e, 6f). The height of the affected roof strata extends up to around 36 m and 55 m, respectively. The three-hinged arch formed in the main roof supports the upper collapsed strata and compacts the lower falling rocks. It also can be seen that the separation between collapsed strata begins to close during face advance.

(4) Main roof last bending After 80 m of face advance, the maximum height of the collapsed strata reaches nearly 70 m. The last periodic bending interval is 15 m. Overall, the height of collapsed strata increases gradually with face advance. The curve in Figure 7 shows this trend.

Analysis of results Coal is mined from right to left in the model. The dynamic movement of overlying strata was recorded by a digital camera during mining. An electronic theodolite was used to keep track of subsidence of strata. These photographs and data illustrate the movement of the overlying rock strata and roof caving and bending.

Roof caving characteristics As the face advances, immediate roof caving initiates, followed by main roof bending, then periodic bending. Figure 6 illustrates this sequence of roof caving.

(1) Immediate roof caving The immediate roof caves in during extraction due to low strength and the large unsupported area. An advance of the face to 25 m initiates the first small-scale caving of the

Figure 5 – Measuring points in the built model

Table II

Equivalent material physicomechanical parameters and model dimension parameters

Mudstone Mid-grain sandstone Mudstone Mid-grain sandstone Fine-grain sandstone Mid-grain sandstone Mudstone Fine-grain sandstone Mudstone Coal Fine-grain sandstone

Strength MPa

Unit weight g/cm3

Model

Field

Model

Field

0.14 0.38 0.14 0.14 0.34 0.38 0.14 0.34 0.14 0.07 0.34

23 65 23 23 58 65 23 58 23 12 58

1.42 1.53 1.42 1.42 1.65 1.53 1.42 1.65 1.42 0.79 1.65

2.42 2.6 2.42 2.42 2.8 2.6 2.42 2.8 2.42 1.35 2.8

The Journal of The Southern African Institute of Mining and Metallurgy

Cumulative Layer Layer thickness thickness weight cm cm kg

Sand weight kg

Lime weight kg

Paste weight kg

Water weight kg

Proportion

101.6 86.6 72.6 57.6 49.8 45.8 41.4 34.5 25.7 11.3 5

65.28 58.80 65.28 32.76 17.16 18.46 30.03 37.68 62.64 27.75 21.42

5.08 4.20 5.08 2.05 1.07 1.31 2.34 2.36 4.87 2.22 1.34

2.18 4.20 2.18 2.05 1.07 1.32 1.00 2.36 2.09 0.56 1.34

5.80 5.38 5.80 3.02 1.54 1.69 2.67 3.39 5.57 2.44 1.93

90:7:3 70:5:5 90:7:3 70:5:5 80:5:5 70:5:5 90:7:3 80:5:5 90:7:3 100:8:2 80:5:5

15 14 15 7.8 4 4.4 6.9 8.8 14.4 6.3 5

72.54 67.20 72.54 36.86 19.30 21.10 33.37 42.40 69.60 30.53 24.10

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Lithology

Investigation into strata behaviour and fractured zone height in a high-seam longwall coal mine

Figure 6 – Model excavation roof caving sequence

After extraction of the face, three vertical zones and three horizontal areas can be distinguished, as shown in Figure 8.

The broken small blocks from the immediate roof are deposited on the floor and are compacted by upper collapsed strata. The main roof is more likely to overhang for a larger distance and break into larger blocks, between which horizontal compressional forces are generated. The collapsed strata form a three-hinged arch shape due to the location of the supporting areas.

Overlying strata subsidence The vertical displacement of the strata can be obtained by the changes in position of the measuring points, which are shown in Figure 9. As can be seen, the measuring points from the upper three measuring lines (A, B, and C, Figure 5) are almost coincident with the horizontal axis, which implies a displacement of almost zero. In contrast, the other strata (lines D to J) undergo significant subsidence, with both the amount of subsidence and the area affected by the subsidence increasing downward. The measurement points in row 5 indicate much greater subsidence than that in row 6, except for the two bottom lines (I and J) in the immediate roof. This is due mainly to the fact that points in row 5 are at the bottom of three-hinged arches. Lines I and J (located in immediate roof) are relatively steady, compared with significant fluctuations found from line E to H.

Figure 7 – Progressive collapsed height of roof strata

Fractured zone height The bulking factor increases with decreasing rock strength (Xia et al., 2014). The bulking factor is 1.1–1.2 if the roof is strong; 1.2–1.3 if the roof is medium strong; and 1.3–1.4 if the roof is weak. The main roof in Wangzhuang coal mine is medium-hard or weak. According to the model, main roof first bending distance (L) is 45 m. The mining thickness (M) is 6.3 m, and the bulking factor (Kρ) is 1.2–1.4. Table III lists the calculated values of (M−0.5×L×sinα)/(Kρ−1) (from Equation [8]) while the rotation angle (α) varies from 3° to 7° and the bulking factor (Kρ) varies from 1.2 to 1.4.

To determine whether the main roof is included in the fractured or caved zone, it is necessary to compare the calculated values in Table III with the thickness of immediate

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Figure 8 – The three vertical and three horizontal zones after extraction of the face

roof (∑h), or the thickness of strata that is potentially included in the caved zone. If Equation [8] is met, a roof of thickness of ∑h is more likely to be incorporated in the fractured zone. If not, this roof is in caved zone.

The Journal of The Southern African Institute of Mining and Metallurgy

Investigation into strata behaviour and fractured zone height in a high-seam longwall coal mine H1=1.1×(14.4+4.4)=20.68 m H2=69.6-20.68=48.92 m The continuous subsidence zone in this physical model is not actually subsiding, while subsidence of the ground surface in the field can be observed. However, the fractured and caved zones in this experiment are clear. The physical model can be improved in future research by, for instance, involving roof sag rates, exploring the influence of different key/competent strata (massive, strong strata above main roof), etc.

Roof fracture investigation

Figure 9 – Strata subsidence

During first main roof bending, ∑h=14.4<25.61 (α=3°, Kρ=1.2, from Table III), so the immediate roof is included in the caved zone. In fact, it is obvious that no horizontal compressional force exists among immediate roof blocks in the model. The immediate roof is therefore assigned to the caved zone.

Then the roof of thickness of ∑h+0.5h is considered.

Since ∑h+0.5h=18.8<20.49 (α=3°, Kρ=1.25), lower main roof is also assigned to the caved zone. In the physical model, we can see that the lower main roof slides and no horizontal compressional force is generated, which verifies the theoretical solution.

Developing this method, ∑h+h=23.2>14.46 (α=5°, Kρ=1.3), hence the upper roof is included in the fractured zone.

It can be seen that upper main roof rotates and a three-hinged arch with horizontal compressional force between the blocks is generated (see Figure 10). The bottom of the upper main roof turns out to be the boundary of the caved and fractured zones.

During periodic main roof bending, the bending interval decreases to 10–15 m, but the bulking factor increases to 1.35 or more. As a result, the main roof is included in the fractured zone.

During the last main roof bending, the bending interval is 15 m, and falling immediate roof blocks are compacted (see Figure 8). It turns out that ∑h+h=23.2<59.07 (α=3°, Kρ=1.1), so the main roof slides and roof blocks fail to generate compressional force. The main roof is included in the caved zone in this weighting.

To summarize, during first main roof weighting, the lower main roof is included in the caved zone but upper main roof is included in the fractured zone; in next two roof periodic weightings, the whole main roof is included in the fractured zone; but in the last main roof bending, the whole main roof is included in the caved zone. Thus the boundary between the caved and the fractured zones changes with face advance. In this study, it is believed that this boundary is the bottom of upper main roof (or top of lower main roof). The heights of the caved and fractured zones are given in Equations [11] and [12].

Boreholes were drilled into the roof in the transportation roadway of Panel 8101, and a borehole camera was used to record the fractures in the roof. Some fractures can be clearly seen (Figure 11).

Unfortunately the length of the boreholes is less than 7 m, thus the plotted fractures in the roof above the transportation roadway are from the caved zone.

Photographs of fractures in the fractured zone could not be obtained. The fractures in different boreholes are mainly horizontal continuous fractures, and rather intensive, indicating that the fractures might be connecting by a horizontal fracture plane in the roof.

Conclusions ➤ Mining equipment, especially shields, enables annual production of high single-pass longwalling to increase to 14 Mt, the mining height to increase to 7 m, and recovery to increase to 96.7% in some longwalls in China. Further improvement in thick-seam coal mining will depend on innovation and the development of equipment ➤ The weak immediate roof is generally included in the caved zone. Breakage of the massive strong roof is classified into rotation and sliding, which determines whether the roof is included in the caved or fractured zone. Main roof sliding causes this roof to be

Table III

Calculated values Kρ/α

3° 5° 7°

1.2

1.25

1.3

1.35

1.4

25.61 21.69 17.79

20.49 17.36 14.23

17.07 14.46 11.86

14.63 12.40 10.17

12.81 10.84 8.89

[11] [12]

where Kρ’ (=1.1) is the ultimate bulking factor. H0 (=69.6 m) is the collapsed height, H1 is the caved zone height, and H2 is the fractured zone height.

As a result, the calculated caved and fractured zone heights shown below are quite close to the results from empirical equations.

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Figure 10 – Boundary between caved and fractured zones VOLUME 115

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H2 = H0 – H1

Investigation into strata behaviour and fractured zone height in a high-seam longwall coal mine incorporated in the caved zone, while main roof rotation incorporates it in the fractured zone, with horizontal forces generated between roof blocks ➤ A theoretical technique has been developed to determine the breakage mode of the roof, and thus the caved and fractured zone heights. The fractured zone height is closely related to mining height (M), immediate roof thickness (∑h), bulking factor (Kρ), main roof strength (RT), main roof thickness (h), main roof weighting interval (L), and overlying strata stress (q). Of these, the bulking factor, mining height, and roof thickness are more relevant ➤ Representative physical modelling displays the three vertical zones and three horizontal areas. The collapsed height of the overlying strata is about 70 m. The main roof first weighting distance is 45 m, and the periodic weighting interval stands at 10–15 m. The maximum subsidence is 62 mm ➤ According to the analytical and physical models, the caved zone height is about 21 m and the fractured zone height is around 49 m, which is close to the results from the empirical equation.

Acknowledgements The authors are grateful for Dr Sam Spearing’s help. This paper is supported by The National Basic Research Program (973 Program, 2013CB227903), National Natural Science Foundation (U1361209), and CUMTB Research Fund for the Doctoral Program (00-800015z683).

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ZHAO, J.L. 2004. Research on new method of full-seam mining for gently inclined thick coal seams. China Coal Industry Publishing House, Beijing (in Chinese).

ZHAO, Y.X., JIANG, Y.D., LV, Y.K., and CUI, Z.J. 2013. Similar simulation experiment of bi-direction loading for floor destruction rules in coal mining above aquifer. Journal of China Coal Society, vol. 38, no. 3. pp. 384–390 (in Chinese). ◆

Figure 11 – Fractures in the roof

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Parametric estimation of capital costs for establishing a coal mine: South Africa case study

by M. Mohutsiwa* and C. Musingwini†

Capital cost estimates are important in decisions on whether a project will be approved, mothballed, or abandoned. In South Africa, junior coal miners do not have extensive databases of historical projects from which to estimate capital costs. The purpose of this paper is to establish formulae that can be used for estimating capital costs of developing coal mines in a coal-producing country, using South Africa as a case study. The costs are estimated to an error of magnitude level of -30% to +50%, which is suitable for a concept study level, using a parametric estimation technique.

The study uses data from completed coal mining projects from selected coal-producing countries. Three formulae are developed and presented for estimating capital costs of underground bord and pillar, surface shovel and truck, and dragline operations.

Keywords coal mining, capital costs, parametric cost estimation.

Introduction Capital cost estimates play a critical role in deciding whether projects will proceed, be delayed, or abandoned. It is therefore important that capital cost estimates are carried out accurately as determined by the estimation guidelines based on the level of estimation conducted (Shafiee and Topal, 2012). The lack of literature on estimation of capital costs in many coal mining countries is adversely affecting junior coal miners, since they do not have an extensive database of historical projects on which to base estimations. In South Africa, the difficulty in estimating capital costs is attributable to lack of publicly available data on completed South African mining projects (Hall, 2013).

Harper (2008) estimated the capital costs of setting up coal mines in Australia, and Dipu (2011) conducted a similar exercise for India, but prior to this study no such investigation had been done for countries such as South Africa. The present investigation uses parametric estimation for estimating capital costs of developing coal mines in a coalproducing country, using South Africa as a case study.

A generic mining cost structure There are two main types of costs for mining projects – operating and capital costs. Figure 1 The Journal of The Southern African Institute of Mining and Metallurgy

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Capital costs Capital costs in mines can be split into start-up (initial) and stay-in-business (working) capital costs. Start-up costs focus mainly on accessing the orebody, infrastructure (mining and beneficiation), equipment, environmental compliance, and licensing costs. For underground mines, accessing the orebody involves sinking shafts or adits, whereas for surface mines it involves the development of boxcuts and removal of initial overburden. It is common practice in the mining industry to hire contractors to carry out the initial work.

Infrastructure covered by start-up capital includes water and electricity connections, offices, workshops, change houses, roads, and employee accommodation. In some instances, small towns and off-mine infrastructure such as rail, roads, and ports need to be developed to accommodate employees and for transportation of coal (Rudenno, 2009). The infrastructure requirements also include mineral processing facilities or plants. The start-up capital expenditure is normally undertaken as quickly as is practically possible in order to move the project into production so that revenues can be generated (Rudenno, 2009). To reduce the costs associated with offmine infrastructure, it is common practice for mining companies to partner with the host government in what is known as publicprivate partnerships (PPPs) to develop such infrastructure (Jourdan, 2008).

Once the mining project has taken off, additional infrastructure, development,

* Standard Bank of South Africa, Johannesburg, South Africa. † School of Mining Engineering, University of Witwatersrand, Johannesburg, South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Aug. 2014 and revised paper received Apr. 2015 AUGUST 2015

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Synopsis

is a schematic representation of the generic cost breakdown for mines. A brief overview of what these costs entail is discussed in the following sections.

Parametric estimation of capital costs for establishing a coal mine

Figure 1 – Generic mining cost structure

equipment overhaul and replacement is financed through stay-in-business capital. Stay-in-business capital does not cover operational cost items that are needed to keep the mine running (Rudenno, 2009). Figure 2 shows a typical distribution of total capital costs for relevant activities in the production cycle in open-pit mines. The typical mining cost breakdown shown in Figure 2 was established by analysing the costs of different mines worldwide, as obtained from Anglo American Thermal Coal and the Raw Material Group’s (RMG) databases.

Operating costs Operating costs focus on the day-to-day running of a mine.

These costs can be divided into fixed and variable costs.

Variable costs vary with the tonnage produced; these costs include fuel, explosives, and electricity costs. Fixed costs, on the other hand, are independent of the tonnage produced, and they include labour costs that are not linked to production. Generally, operating costs are expressed in total costs per ton of ore mined; for example, South African coal mining costs are quoted as rands per ton of coal produced.

Figure 3 shows typical operating cost splits for relevant activities in the production cycle in open-pit mines. It can be seen that that hauling typically accounts for a significant proportion of the operating costs. The lowest cost activity in the production cycle is drilling.

Figure 2 – Typical distribution of total capital costs in open-pit mines

Comparison of coal mining sectors by country Mining costs differ from country to country due to factors such as general geology of the orebodies, infrastructure, labour productivity, and the level of economic development.

According to the World Coal Institute, there are significant similarities between the South African and Indian coal mining sectors. The comparison was based mainly on coal geology, mining methods used, and labour productivity.

Table I shows the relative contributions of room-and-pillar and longwall mining to underground coal extraction in seven coal-producing countries.

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Figure 3 – Distribution of costs for main activities in the production cycle in open-pit mines (Source: Bagherpour, 2007) The Journal of The Southern African Institute of Mining and Metallurgy

Parametric estimation of capital costs for establishing a coal mine Background to cost estimation

Table I

Relative contribution of room-and-pillar and longwall mining to underground coal production in selected countries (Source: World Coal Institute) Country

Percentage of underground production Room and pillar Longwall

India South Africa USA Australia China Germany United Kingdom

91 91 55 28 No information 0 <1

5 9 45 72 20 100 99

The similarities between South African and Indian coal mining methods play a significant role in the capital and operating cost structures. South African room-and-pillar mining operations are at shallow depths (generally less than 100 m) while the average range in India is about 300 to 350 m (Xie, 2008).

Mining costs in Indian mines are about 35% higher than costs in leading coal-exporting countries such as Australia, Indonesia, and South Africa (Santra and Bangaria, 2014).

The higher mining costs in India are driven mainly by low labour productivity. In 2013 India forecast a production increase of 3%, compared to only a 2% increase in South Africa.

Cost estimation can be defined as a predictive process used to quantify cost and price the resources required by the scope of an investment option, activity, or project (Leo and Knotowicz, 2005). Cost estimation can also be regarded as a process used to predict the uncertainty of future costs; in this context, the goal of cost estimating is to minimize the uncertainty of the estimate given the level of scope and definition (Leo and Knotowicz, 2005). The cost estimating process is generally applied during each phase of the project’s life cycle and whenever the project scope is re-defined, modified, or refined. As the level of scope definition increases, the estimating methods used become more definitive and produce estimates with increasingly narrower probabilistic cost distributions.

The two fundamental approaches to estimating costs are the top-down and bottom-up approaches. The top-down approach uses historical data from similar projects. It is best used when alternatives are still being developed and refined (Sullivan, Wicks, and Koelling, 2012). The bottom-up approach, on the other hand, is more detailed and works best when the detail concerning the desired output (product or service) has been defined and clarified (Sullivan, Wicks, and Koelling, 2012). In general, the bottom-up approach is more detailed than the top-down approach. Figure 4 shows the simplified cost estimation processes for the two approaches.

The results of cost estimations are used for a variety of purposes, including: ➤ Planning the appropriate funding strategy

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Figure 4 – Simplified cost estimation processes for bottom-up and top down approaches (Source: Novellaqalive2, n.d.)

Parametric estimation of capital costs for establishing a coal mine ➤ Enabling decisions on the viability of feasibility studies ➤ Project analyses and evaluations in the research and development phase ➤ Evaluating alternative investments ➤ Serving as a basis for cost control activity during job execution.

Cost estimation in mining According to Hall (2013), there is a lack of general estimates of capital costs for establishing coal mines in South Africa. In a 2008 regression analysis study, the Australian Average Capital Costs (ACC) for developing an open-pit mine were estimated as A$53 million plus A$33 million per 1 Mt/a of coal produced and treatment capacity, while the ACC for an Australian underground mine was estimated as A$37 million plus A$68 million per 1 Mt/a of production and treatment capacity (Harper, 2008).

In 2011 costs, the capital costs of surface coal mines in India were estimated to range from US$31.7–44.3 per ton of rated capacity. This means that a 1 Mt/a capacity mine can be expected to cost approximately US$44.3 million (Dipu, 2011).

The above estimation is based on a stripping ratio of 4:1, and appropriate adjustments can be made to account for higher or lower stripping ratios. For underground mining, the estimates range is between US$40.7 and US$59.1 per ton of rated capacity. This means that a 1 Mt/a rated Indian coal mine will require an estimated capital cost of US$59.1 million (Dipu, 2011). The underground estimates are applicable to shallow mines (less than 150 m depth) operated with semimechanized bord-and-pillar mining methods.

Brief description of the parametric cost estimation methodology Cost estimation was conducted using a parametric estimation model derived from parametric estimating techniques. A parametric estimation model is a mathematical representation of cost relationships that provides a logical and predictable correlation between the physical or functional characteristics of a project and its resultant cost (ISPA, 2008). According to Dysert (2005), parametric cost estimating models are useful tools for preparing early conceptual estimates where there is little technical information available to support the use of more detailed estimating methods.

In this study, actual project costs were used as dependent variables, whereas factors such as the main mine capacity, life of the mine, and stripping ratio were used as independent variables. Relationships between dependent and independent variables from historical data on global coal mines were used to estimate costs for future coal mining projects within a coalmining country. The model runs from an MS Excel® spreadsheet and capital costs are estimated to an order-ofmagnitude level, which is appropriate for use at a concept study level. With all the cost models, data is normalized to ensure consistency.

The first step in developing the model was determining the scope. The cost drivers for coal mines were identified to be studies, project management, site preparation, infrastructure, equipment, stores, and sundry. The abovementioned cost drivers were not considered individually when estimating the total capital cost, due to lack of data.

Figure 5 illustrates the build-up of the database to account for cost differences embedded in different mining methods.

Figure 5 – Database build-up to account for different coal mining methods

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Parametric estimation of capital costs for establishing a coal mine Table II

Surface operations employing draglines for waste removal

Amlohri Bina Dudhichua Jayant Khadia Nigahi Vista Washpool ATCOM OP Middelburg New Largo Isibonelo

Country

Actual project cost (US$ million)

India India India India India India Canada Australia South Africa South Africa South Africa South Africa

245.1 36.2 70 227.2 319.1 451.4 444.5 354.9 407 975 530 68

Data was obtained from the Anglo American Thermal Coal and Raw Materials Group (RMG) databases. The cost data looks into actual costs, designed production capacities, stripping ratios, and run-of-mine (ROM) production.

As shown in Figure 5, data was first segmented into surface and underground operations. The rationale behind splitting the database in this way is to capture the influence of different mining methods on capital costs. Data for surface mining operations is further split into dragline operations and truck-and-shovel operations. Dragline operations consider a combination of truck-and-shovel and dragline as the main waste removal equipment. Underground operations are divided into room-and-pillar and longwall operations. Owing to the lack of longwall mining data-set for South Africa, and no immediate plans to build such operations, capital costs were not estimated for this type of mining method.

Dragline operations Dragline operations include mines that use a combination of dragline and truck-and-shovel for primary waste removal.

Table II shows 12 global operations used in this study to develop a formula for estimating the capital costs of establishing surface coal mines employing draglines for waste removal in selected coal-mining countries, including South Africa. The actual project costs were normalized (adjusted to 2013 fiscal year) by using mining indices.

Table II shows that draglines are considered where the LOM exceeds 10 years. This is in part attributable to the high capital requirement for dragline equipment. Variations in production capacity in Table II show that draglines can be used for operations with less than 5 Mt/a production, probably if the stripping ratio is high and significant waste removal may be required.

As shown in Figure 6, mines with a higher rated capacity attract lower capital costs than mines with lower rated capacity. The capital cost and rated capacity relationship shown in Figure 6 can be attributed to the concept of economies of scale.

Table III shows the regression results used to develop a formula for surface coal mines using draglines. Based on the regression, the R2 value is equal to 0.9189, thus 91.89% of the variation in the data about the average is explained.

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Mine life (years)

Capacity (Mt/a)

20 12 20 20 20 20 20 16 20 20 20 15

Stripping ratio

4.0 6.0 10.0 10.0 4.0 15.0 6.0 7.0 2.4 17.0 15.0 4.1

7.17 7.26 7.07 7.07 7.11 7.26 5.10 6.50 4.0 2.0 6.0 5.0

The best-fit line for the above data is given by: [1]

Truck-and-shovel operations Truck-and-shovel is a commonly used method for surface coal mining. For this study, the operations were not standardized for either truck or shovel size. The actual project costs were normalized (adjusted to 2013 fiscal year) by using mining indices. The reason for using a global sample is the lack of sufficient data on completed operations in South Africa. Table IV lists the truck-and-shovel operations used in this study.

Figure 7 shows the relationship between capacity and actual project cost for truck-and-shovel operations.

The operations used are based in South Africa and India.

The design capacity of the operations ranges from 1.0 to 25.0 Mt/a (ROM). The variation in the design capacity is a clear indication of the flexibility of truck-and-shovel operations.

Table V shows the regression results used to develop a formula for surface coal mines using truck-and-shovel. Based on the regression in Table V, the R2 value is equal to 0.9882, meaning that 98.82% of the variation within the truck-andshovel date used is explained.

From Table V, the capital cost of establishing surface coal mines that use truck-and-shovel was determined to be:

Figure 6 â&#x20AC;&#x201C; Relationship between the actual project cost and mine capacity VOLUME 115

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Parametric estimation of capital costs for establishing a coal mine Table III

Regression results for surface dragline method

Note: df (degrees of freedom); SS (sum of squares); MS (mean square); t-Stat (Student’s t-statistic)

Table IV

Truck-and-shovel operations Operation

Country

Balaram Baround Top Seam Basundhara Gevra Hingula Kaniha OP Kulda Kusmunda Moabsvelden Middelburg O/C

India India India India India India India India South Africa South Africa

Project actual cost (US$ million) 85.4 42.7 38.6 293.2 74.9 100.2 293.1 98.6 31.8 975

Mine life (years)

Capacity (Mt/ya)

20 20 20 20 20 20 20 20 15 20

8.0 1.0 2.4 25.0 8.0 10.0 10.0 3.5 3.0 20.0

Stripping ratio

2.76 2.92 2.76 2.99 2.76 2.76 3.60 2.99 1.95 5.00

[2]

Bord-and-pillar mining method

Figure 7 – Relationship between the actual project cost and mine capacity

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The bord-and-pillar mining method as referred to in this report is limited to mechanized coal mining. The capital cost differences between traditional and conventional mining methods made it impractical to combine the cost data of two test mining methods. Table VI shows the operations used in the regression analysis.

Figure 8 shows the relationship between capacity and actual costs of establishing underground bord-and-pillar coal mines.

Table VII shows the statistical output of the regression analysis of underground bord-and-pillar operations. Based on the results, 90.58% of the variance between actual costs, mine life, and capacity is explained.

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Parametric estimation of capital costs for establishing a coal mine Table V

Regression results for surface truck and shovel mining method

Table VI

Bord-and-pillar underground operations Operation

Country

Balaram Baround Top Seam Basundhara Gevra Hingula Kaniha OP Kulda Kusmunda Moabsvelden Middelburg O/C

India India India India India India India India South Africa South Africa

Project actual cost (US$ million)

Mine life (years)

85.4 42.7 38.6 293.2 74.9 100.2 293.1 98.6 31.8 975

Capacity (Mt/a)

20 20 20 20 20 20 20 20 15 20

Stripping ratio

8.0 1.0 2.4 25.0 8.0 10.0 10.0 3.5 3.0 20.0

2.76 2.92 2.76 2.99 2.76 2.76 3.60 2.99 1.95 5.00

[3]

Results

Once the coefficients were determined, Equation [3] for estimating capital costs was developed and used on the initial cost data to calculate the error between the actual and estimated costs.

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Figure 8 – Relationship between the actual project cost and mine capacity

The formulae derived from regression analyses were used to estimate capital costs, which were later compared with the actual costs. The results are presented in Tables VIII–X. Using Equation [1], the error between actual project costs and estimated project costs range from zero to 489%. Three of the Indian operation (not shown in Table VIII) were classified as outliers and removed from the regression.

Research conducted on the Isibonelo operation to further understand the drivers behind the high percentage error revealed that some of the major items of equipment used to establish the operation were not purchased, but were moved from existing operations. Based on the calculated error, the formula can be used to estimate capital costs of establishing

Parametric estimation of capital costs for establishing a coal mine Table VII

Regression results for underground bord and pillar mining method

Table VIII

Actual and estimated cost comparison for dragline operations Operation

Country

Amlohri Khadia Nigahi Vista Washpool ATCOM OP Middelburg New Largo Isibonelo

India India India Canada Australia South Africa South Africa South Africa South Africa

Actual project cost (US$ million) 245.1 319.1 451.4 444.5 354.9 407 975 530 68

coal mines in a coal-producing country to an order-ofmagnitude level.

From Table IX it can be seen that Basundhara coal mine in India is an outlier. The calculated error between actual and estimated projects costs range from -23% to 27%. Based on the definition and error limits of ’error of magnitude’ estimation level, it can be concluded that Equation [2] can be used to estimate the cost of establishing surface truck-andshovel operations to an error-of-magnitude level.

From Table X, it can be seen that the calculated error ranges from -52% to 27%. Irenedale mine cost data is regarded as an outlier and was eliminated from further analyses. Some of the factors that influenced the Irenedale cost data are the deeper shafts compared to other operations and the establishment of a surface overland conveyor.

Based on the parameters of order-of-magnitude estimation level, it can be concluded that Equation [3] can be used to to estimate the capital costs of establishing mechanized underground coal mines in a coal-producing country to an order–of-magnitude level.

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279.74 283.62 451.21 445.52 354.9 458.47 822.78 532.50 400.83

Percentage error

14% -11% 0% 0% 0% 13% -16% 0% 489%

Discussion and conclusion The scarcity of capital cost data on the lowest cost-activity levels is a challenge in the mining industry. Most databases only provide data at a high level (total capital costs) without any detailed cost breakdown. The study is limited to countries where off-mine infrastructure has been established to support the development and construction of coal mines.

The study is also limited to areas where skills related to mine development are available.

The three formulae developed in this study can be used to estimate capital costs of establishing coal mines in a coalproducing country to an error-of-magnitude level of -30% to +50% as illustrated by Tables VIII to X. This level of estimation is appropriate at the conceptual study stage. The correlation between capital costs required for establishing coal mines in South Africa and India is quite significant and clearly visible from the resultant errors calculated between actual and estimated costs. No formula could be developed for estimating capital costs of establishing underground longwall mines due to the lack of actual cost data from such mines in both South Africa and India.

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Parametric estimation of capital costs for establishing a coal mine Table IX

Actual and estimated cost comparison for truck-and-shovel operations Operation

Country

Balaram Baround Top Seam Basundhara Gevra Hingula Kaniha OP Kulda Kusmunda Moabsvelden Middelburg O/C

India India India India India India India India South Africa South Africa

Actual project cost (US$ million) 85.4 42.7 38.6 293.2 74.9 100.2 293.1 98.6 31.8 975

Estimated project cost (US$ million)

Percentage error (%)

65.36 53.74 12.37 304.74 65.36 84.29 371.02 101.29 31.80 943.53

-23% 26% -68% 4% -13% -16% 27% 3% 0% -3%

Table X

Actual and estimated cost comparison for bord-and-pillar operations Operation

Country

Irenedale Mbila Anthracite Penumbra Mooiplaats U/G Tumelo U/G Twistdraai U/G

South Africa South Africa South Africa South Africa South Africa South Africa

Actual project cost (US$ million)

In conclusion, this study was able to establish formulae that can be used in the early stages of coal mining projects to estimate the costs of establishing surface and underground coal mines in a coal-producing country to an order-ofmagnitude estimation level.

32.1 88 36.4 141.5 126.8 215.4

Estimated project cost (US$ million)

Percentage error (%)

15.31 90.96 46.27 154.36 142.08 191.21

-52% 3% 27% 9% 12% -11%

Jourdan, P. 2008. Plan of Action for African Acceleration of Industrialization— Promoting Resource-Based Industrialization: a Way Forward. Mimeo, Pretoria,

Leo, D.W. and Knotowicz, S.W. 2005. Project Estimate Reviews. AACE International Transactions. EST 05. Morgantown, WV.

Acknowledgement The work reported in this paper is part of an MSc research study at the University of the Witwatersrand

Novellaqalive2. Not dated. Cost estimation and indirect cost allocation http://novellaqalive2.mhhe.com/sites/dl/free/007096310x/591328/ch18_ CostEstimation.pdf [Accessed 28 May 2014].

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Rudenno, V. 2009. The Mining Valuation Handbook: Mining and Energy Valuation for Investors and Management. 3rd edn. Milton, Queensland, Australia.

mining operations, PhD thesis, Department of Energy and Management, Technical University of Clausthal.

Dipu, D. 2013. Capital costs of Indian coal mining project – an analyst view.

Santra, S. and Bangaria, N. 2014. Labour productivity in coal mining sector in India: with a special focus on major coal mining states. Researchjournali’s Journal of Human Resource, vol. 2, no. 1, January.

http://www.coalspot.com/news/1888/capital-costs-of-indian-coal-miningproject/1 [Accessed 28 May 2014].

Dysert, L.R., 2005. So you think you’re an estimator? AACE International Transaction, EST.01.

Hall, I. 2013. Anglo American Thermal Coal. Personal communication.

Harper, P. 2008. What does a feasibility study cost? International Mining,

Shafiee, S. and Topal, E. 2012. New approach for estimating total mining costs in surface coal mines. Transactions of the Institute of Materials, Minerals and Mining. Section A: Mining Technology, vol. 121, no. (3. pp. 109-116.

Sullivan, W., Wicks, E., and Koelling, C. 2012. Engineering Economy. 15th edn. Prentice Hall.

April 2008, pp 60- 61.

Estimating Handbook, 4th edn. Vienna, VA, USA The Journal of The Southern African Institute of Mining and Metallurgy

Xie, G. 2008. Influence of thickness to the mechanical characteristics of gateways at FMTC faces. Journal of Coal Science Engineering (China), vol.

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International Society of Parametric Analysts (ISPA). 2008. Parametric

INTERNATIONAL CHAIR Dr. Raj K. Singhal (singhal@shaw.ca) HONORARY CHAIR Mr. Mike Teke CHAIRMAN MPES 2015 Prof. Cuthbert Musingwini

3 ECSA POINTS ESSES: KEYNOTE ADDR

R.

M. Teke N. Froneman G. Lane n Webber-Youngma

LOCAL ORGANISING COMMITTEE Dr Bekir Genc Dr Steven Rupprecht Mr Alastair Macfarlane Mr Kelello Chabedi Mr Jannie Maritz Mr Godknows Njowa Mr Mike Woodhall Prof. Jim Porter Mr Alex Bals Dr Andre Dougall Dr Gordon L. Smith Dr Craig Smith Mr I. Wermuth Mr C. Birch Mr G. Lane CO-CHAIRS Professor Monika Hardygóra Prof. Carsten Drebenstedt Prof. Uday Kumar Prof. Kostas Fytas Prof. Petr.Sklenicka INTERNATIONAL ORGANISING COMMITTEE Prof. Hani Mitri Dr Nuray Demirel Dr Marilena Cardu Dr Fiona Mavroudis Dr Meimei Zhang Prof. Ge Hao Prof. Celal Karpuz Prof. Liu Mingju Dr Mohan Yellishetty Prof. Hideki Shimada Dr Gento Mogi Dr Vera Muzgina Dr Morteza Osanloo Dr Juri- Rivaldo Pastarus Prof. Hakan Schunnesson Ms. M. Singhal Dr Eleonora. Widzyk-Capehart Prof. Antonio Nieto Prof. Michael A. Zhuravkov Prof. Sukumar Bandopadhaya Prof. Ferri Hassani Dr Noune.Melkoumian Dr Joerg Benndorf Prof. Giorgio Massacci Dr Maria Menegaki Prof. Svetlana V.Yeffremova Prof. Pivnyak Gennadiy Prof. Vladimir Kebo Dr Marie Vrbova

23rd INTERNATIONAL SYMPOSIUM on MINE PLANNING & EQUIPMENT SELECTION

Smart Innovation in Mining 8 November 2015—Cocktail Function 9–11 November 2015—Conference 12 November 2015—Tours and Technical Visits Sandton Convention Centre, Johannesburg, South Africa BACKGROUND The Southern African Institute of Mining and Metallurgy (SAIMM) will be hosting the 23rd International Symposium on Mine Planning and Equipment Selection (MPES) for the first time in South Africa in 2015.

The symposium has successfully been organized annually for the past 25 years with intermittent 1-year recess as necessary. In 2013 the symposium was held in Germany. It will be in recess in 2014. MPES has previously been held in Turkey, Greece, Canada, Kazakhstan, Australia, Czech Republic, Brazil, India, China, Ukraine, Poland and Italy. Other venues for future MPES are Czech Republic, Sweden, and Australia.

OBJECTIVES The key objective of MPES is to provide a platform for researchers from academic institutions, professionals from mining companies, practitioners from consulting companies, equipment suppliers (OEMs) and software providers to share the latest global developments in mine planning and equipment selection across all commodities for the benefit of the mining industry in improving efficiencies and safety.

An additional objective of MPES 2015 is to encourage MSc and PhD students to showcase their research in a ‘Young Authors Category’ to foster creation of the next generation of MPES participants.

MAJOR THEMES TO BE COVERED ALONG THE VALUE CHAIN • • • • • • • • • • • • • • •

Data Collection and Modelling: State of the Art Practices Mineral Resource and Mineral Reserve Estimation and reporting Economic and Technical Feasibility Studies, Mine Development Case Studies Design, Planning and Optimization of Surface and Underground Mines Transition from surface to underground mining Rock Mechanics and Geotechnical Applications Mining Equipment: Selection, Operation, Control, Monitoring and Optimization Mechanization and Automation of Mining Processes Application of Information Technology Short interval/planning and control Resource-to-Market: Reconciliation and Optimization Productivity and Competitiveness of Mining Operations Sustainability: Improving Health, Safety and Environmental Practice and Performance Mine Closure and Rehabilitation in mine planning Young Authors Category (MSc/PhD Students below 35 years).

Confirmed Technical Visits: Please note that a maximum of 20 delegates can be accommodated on each of these Technical Visits • Zibulo Colliery (Underground Coal)

• Grootegeluk Mine (Open Pit Multi-Seam)

• Bathopele Mine (Mechanised Platinum)

• Tau Tona Mine (Deep Level Gold Mine)

• University of Pretoria Virtual Reality Laboratory

Sponsor:

EXHIBITION/SPONSORSHIP Sponsorship opportunities are available. Companies wishing to sponsor or exhibit should contact the Conference co-ordinator.

SHELL PECTEN

CONFERENCE ANNOUNCEMENT

For further information contact, please contact: Dr Raj Singhal, E-mail: singhal@shaw.ca or Conference Organisers: The Southern African Institute of Mining and Metallurgy (SAIMM) · E-mail: raymond@saimm.co.za · Official Website: http://www.saimm.co.za

INTERNATIONAL ACTIVITIES 28 September-2 October 2015 — WorldGold Conference 2015 Misty Hills Country Hotel and Conference Centre, Cradle of Humankind, Gauteng, South Africa Contact: Camielah Jardine Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156, E-mail: camielah@saimm.co.za Website: http://www.saimm.co.za 12–14 October 2015 — Slope Stability 2015: International Symposium on slope stability in open pit mining and civil engineering In association with the Surface Blasting School 15–16 October 2015 Cape Town Convention Centre, Cape Town Contact: Raymond van der Berg Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: raymond@saimm.co.za Website: http://www.saimm.co.za 20 October 2015 — 13th Annual Southern African Student Colloquium Mintek, Randburg, Johannesburg Contact: Yolanda Ramokgadi Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: yolanda@saimm.co.za Website: http://www.saimm.co.za 21–22 October 2015 — Young Professionals 2015 Conference Making your own way in the minerals industry Mintek, Randburg, Johannesburg Contact: Camielah Jardine Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: camielah@saimm.co.za Website: http://www.saimm.co.za 28–30 October 2015 — AMI: Nuclear Materials Development Network Conference Nelson Mandela Metropolitan University, North Campus Conference Centre, Port Elizabeth Contact: Raymond van der Berg Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: raymond@saimm.co.za Website: http://www.saimm.co.za 8–12 November 2015 — MPES 2015: Twenty Third International Symposium on Mine Planning & Equipment Selection Sandton Convention Centre, Johannesburg, South Africa Contact: Raj Singhal E-mail: singhal@shaw.ca or E-mail: raymond@saimm.co.za Website: http://www.saimm.co.za

2016 14–17 March 2016 — Diamonds still Sparkle 2016 Conference Gaborone Contact: Yolanda Ramokgadi

The Journal of The Southern African Institute of Mining and Metallurgy

Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: yolanda@saimm.co.za Website: http://www.saimm.co.za 13–14 April 2016 — Mine to Market Conference 2016 South Africa Contact: Yolanda Ramokgadi Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: yolanda@saimm.co.za Website: http://www.saimm.co.za 17–18 May 2016 — The SAMREC/SAMVAL Companion Volume Conference Johannesburg Contact: Raymond van der Berg Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: raymond@saimm.co.za Website: http://www.saimm.co.za 21–28 May 2016 — ALTA 2016 Perth, Western Australia Contact: Allison Taylor Tel: +61 (0) 411 692 442 E-mail: allisontaylor@altamet.com.au Website: http://www.altamet.com.au May 2016 — PASTE 2016 International Seminar on Paste and Thickened Tailings Kwa-Zulu Natal, South Africa Contact: Raymond van der Berg Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: raymond@saimm.co.za Website: http://www.saimm.co.za 9–10 June 2016 — 1st International Conference on Solids Handling and Processing A Mineral Processing Perspective South Africa Contact: Raymond van der Berg Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: raymond@saimm.co.za Website: http://www.saimm.co.za 1–3 August 2016 — Hydrometallurgy Conference 2016 ‘Sustainability and the Environment’ in collaboration with MinProc and the Western Cape Branch Cape Town Contact: Raymond van der Berg Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: raymond@saimm.co.za Website: http://www.saimm.co.za 11–14 August 2016 — The Tenth International Heavy Minerals Conference ‘Expanding the horizon’ Sun City, South Africa Contact: Camielah Jardine Tel: +27 11 834-1273/7, Fax: +27 11 838-5923/833-8156 E-mail: camielah@saimm.co.za Website: http://www.saimm.co.za

AUGUST 2015

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2015

Company Affiliates The following organizations have been admitted to the Institute as Company Affiliates AECOM SA (Pty) Ltd

Elbroc Mining Products (Pty) Ltd

Namakwa Sands (Pty) Ltd

AEL Mining Services Limited

Engineering and Project Company Ltd

New Concept Mining (Pty) Limited

Air Liquide (PTY) Ltd

eThekwini Municipality

Northam Platinum Ltd - Zondereinde

AMEC Mining and Metals

Exxaro Coal (Pty) Ltd

Osborn Engineered Products SA (Pty) Ltd

AMIRA International Africa (Pty) Ltd

Exxaro Resources Limited

ANDRITZ Delkor(Pty) Ltd

Fasken Martineau

Anglo Operations Ltd

FLSmidth Minerals (Pty) Ltd

Anglo Platinum Management Services (Pty) Ltd

Fluor Daniel SA (Pty) Ltd

Anglogold Ashanti Ltd Atlas Copco Holdings South Africa (Pty) Limited

Outotec (RSA) (Proprietary) Limited PANalytical (Pty) Ltd Paterson and Cooke Consulting Engineers (Pty) Ltd

Franki Africa (Pty) Ltd Johannesburg

Polysius A Division Of Thyssenkrupp Industrial Solutions (Pty) Ltd

Fraser Alexander Group

Precious Metals Refiners Rand Refinery Limited

Glencore

Aurecon South Africa (Pty) Ltd

Redpath Mining (South Africa) (Pty) Ltd

Goba (Pty) Ltd

Rosond (Pty) Ltd

Aveng Moolmans (Pty) Ltd

Hall Core Drilling (Pty) Ltd

Axis House (Pty) Ltd

Hatch (Pty) Ltd

Bafokeng Rasimone Platinum Mine

Herrenknecht AG

Barloworld Equipment -Mining

HPE Hydro Power Equipment (Pty) Ltd

Rustenburg Platinum Mines Limited

BASF Holdings SA (Pty) Ltd

Impala Platinum Limited

SAIEG

Bateman Minerals and Metals (Pty) Ltd

IMS Engineering (Pty) Ltd

Salene Mining (Pty) Ltd

BCL Limited

JENNMAR South Africa

Becker Mining (Pty) Ltd

Joy Global Inc. (Africa)

Sandvik Mining and Construction Delmas (Pty) Ltd

BedRock Mining Support (Pty) Ltd

Leco Africa (Pty) Limited

Sandvik Mining and Construction RSA(Pty) Ltd

Bell Equipment Company (Pty) Ltd

Longyear South Africa (Pty) Ltd

SANIRE

Blue Cube Systems (Pty) Ltd

Lonmin Plc

Sasol Mining(Pty) Ltd

Bluhm Burton Engineering (Pty) Ltd

Ludowici Africa

Scanmin Africa (Pty) Ltd

Blyvooruitzicht Gold Mining Company Ltd

Lull Storm Trading (PTY)Ltd T/A Wekaba Engineering

Sebilo Resources (Pty) Ltd

Magnetech (Pty) Ltd

Senmin International (Pty) Ltd

Magotteaux(PTY) LTD

Shaft Sinkers (Pty) Limited

MBE Minerals SA Pty Ltd

Sibanye Gold (Pty) Ltd

MCC Contracts (Pty) Ltd

Smec SA

MDM Technical Africa (Pty) Ltd

SMS Siemag South Africa (Pty) Ltd

Metalock Industrial Services Africa (Pty)Ltd

SNC Lavalin (Pty) Ltd

Metorex Limited

Sound Mining Solutions (Pty) Ltd

BSC Resources CAE Mining (Pty) Limited Caledonia Mining Corporation CDM Group CGG Services SA Chamber of Mines Concor Mining Concor Technicrete Council for Geoscience Library CSIR-Natural Resources and the Environment

Royal Bafokeng Platinum Roymec Tecvhnologies (Pty) Ltd Runge Pincock Minarco Limited

Metso Minerals (South Africa) (Pty) Ltd Minerals Operations Executive (Pty) Ltd MineRP Holding (Pty) Ltd

SENET

South 32 SRK Consulting SA (Pty) Ltd Technology Innovation Agency Time Mining and Processing (Pty) Ltd

Department of Water Affairs and Forestry

Mintek

Deutsche Securities (Pty) Ltd

MIP Process Technologies

Digby Wells and Associates

Modular Mining Systems Africa (Pty) Ltd

Umgeni Water

Downer EDI Mining

MSA Group (Pty) Ltd

VBKOM Consulting Engineers

DRA Mineral Projects (Pty) Ltd

Multotec (Pty) Ltd

Webber Wentzel

DTP Mining

Murray and Roberts Cementation

Weir Minerals Africa

Duraset

Nalco Africa (Pty) Ltd

WorleyParsons (Pty) Ltd

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AUGUST 2015

Tomra Sorting Solutions Mining (Pty) Ltd Ukwazi Mining Solutions (Pty) Ltd

The Journal of The Southern African Institute of Mining and Metallurgy

Forthcoming SAIMM events...

IP PONSORSH EXHIBITS/S ng to sponsor ishi Companies w ese t at any of th bi hi ex or d/ an e th t ac cont events should rdinator -o co ce conferen ssible as soon as po

SAIMM DIARY 2015 or the past 120 years, the Southern African Institute of Mining and Metallurgy, has promoted technical excellence in the minerals industry. We strive to continuously stay at the cutting edge of new developments in the mining and metallurgy industry. The SAIMM acts as the corporate voice for the mining and metallurgy industry in the South African economy. We actively encourage contact and networking between members and the strengthening of ties. The SAIMM offers a variety of conferences that are designed to bring you technical knowledge and information of interest for the good of the industry. Here is a glimpse of the events we have lined up for 2015. Visit our website for more information.

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◆ CONFERENCE World Gold Conference 2015 28 September–2 October 2015, Misty Hills Country Hotel and Conference Centre, Cradle of Humankind, Muldersdrift ◆ SYMPOSIUM International Symposium on slope stability in open pit mining and civil engineering 12–14– October 2015 In association with the Surface Blasting School 15–16 October 2015, Cape Town Convention Centre, Cape Town ◆ COLLOQUIUM 13th Annual Southern African Student Colloquim 2015 20 October 2015, Mintek, Randburg, Johannesburg ◆ CONFERENCE Young Professionals 2015 Conference 21–22 October 2015, Mintek, Randburg, Johannesburg ◆ CONFERENCE AMI: Nuclear Materials Development Network Conference 28–30 October 2015, Nelson Mandela Metropolitan University, North Campus Conference Centre, Port Elizabeth ◆ SYMPOSIUM MPES 2015: Twenty Third International Symposium on Mine Planning & Equipment Selection 8–12 November 2015, Sandton Convention Centre, Johannesburg, South Africa

2016 ◆ CONFERENCE Diamonds still Sparkle 2016 Conference 14–17 March 2016, Gabarone ◆ CONFERENCE Mine to Market Conference 2016 13–14 April 2016, South Africa ◆ CONFERENCE The SAMREC/SAMVAL Companion Volume Conference 17–18 May 2016, Johannesburg

For further information contact: Conferencing, SAIMM P O Box 61127, Marshalltown 2107 Tel: (011) 834-1273/7 Fax: (011) 833-8156 or (011) 838-5923 E-mail: raymond@saimm.co.za

◆ SEMINAR PASTE 2016 International Seminar on Paste and Thickened Tailings May 2016, Kwa-Zulu Natal, South Africa ◆ CONFERENCE 1st International Conference on Solids Handling and Processing A Mineral Processing Perspective 9–10 June 2016, South Africa

Website: http://www.saimm.co.za

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