Saimm 201504 apr

VOLUME 115

NO. 4

APRIL 2015

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The Southern African Institute of Mining and Metallurgy PAST PRESIDENTS OFFICE BEARERS AND COUNCIL FOR THE 2014/2015 SESSION Honorary President Mike Teke President, Chamber of Mines of South Africa Honorary Vice-Presidents Ngoako Ramatlhodi Minister of Mineral Resources, South Africa Rob Davies Minister of Trade and Industry, South Africa Naledi Pando 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 DRC

S. Maleba

Johannesburg

I. Ashmole

Namibia

N. Namate

Pretoria

N. Naude

Western Cape

C. Dorfling

Zambia

H. Zimba

Zimbabwe

E. Matinde

Zululand

C. Mienie

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

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*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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*

*

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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

VOLUME 115

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.

APRIL 2015

Contents Journal Comment

by H.R. Phillips . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

President’s Corner

by J.L. Porter . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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Special Articles South African National Committee on Tunelling Young Members Group –SANCOT - YMG

by L. Nene. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Handover of model stope to Wits School of Mining Engineering

by S. Braham. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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New head of Wits mining school announced

by S. Braham. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

Wits-SRK link boosts rock engineering skills

by S. Braham. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

The Secretariat The Southern African Institute of Mining and Metallurgy ISSN 2225-6253 (print)

NO. 4

ix x

STUDENT EDITION Re-aligning the cutting sequence with general support work and drafting a support sequence at Simunye Shaft

by K. Lombard. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

251

Mining through areas affected by abnormal stress conditions at Syferfontein Colliery

by C. Legote. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

265

A critical evaluation of the water reticulation system at Vlaklaagte Shaft, Goedehoop Colliery

by R. Lombard. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

271

Optimization of shuttle car utilization at an underground coal mine

by P.R. Segopolo . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

285

Explosives utilization at a Witwatersrand gold mine

by M. Gaula. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

297

Critical investigation into the problems surrounding pillar holing operations

by J.P. Labuschagne, H. Yilmaz, and L. Mpolokeng. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

307

LHD optimization at an underground chromite mine

by W. Mbhalati . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

313

The viability of using the Witwatersrand gold mine tailings for brickmaking

by M. Malatse and S. Ndlovu. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

321

Evaluation of some optimum moisture and binder conditions for coal fines briquetting

by P. Venter and N. Naude . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

329

Air drying of fine coal in a fluidized bed

by M. Le Roux, Q.P. Campbell, M.J. van Rensburg, E.S. Peters, and C. Stiglingh. . . . . . . . . . . .

335

International Advisory Board VOLUME 115

NO. 4

APRIL 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 H. Potgieter, Manchester Metropolitan University, United Kingdom E. Topal, Curtin University, Australia

The Journal of The Southern African Institute of Mining and Metallurgy

APRIL 2015

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

Journal Comment s part of its commitment to supporting young professionals entering the mining and metallurgical industries, the SAIMM holds a Student Colloquium every year. The ten papers in this edition of the Journal are based on presentations made at that event by students and recent graduates in mining engineering, metallurgy, and minerals processing. The opportunity to present their final year research projects is restricted to those students from each institution achieving the best results for this subject, and the papers presented here have been further selected following scrutiny by a panel of senior professionals, acting as judges at the Colloquium.

This year there is an emphasis on coal, with four of the six mining papers based on vacation work at coal mines, while two of the four mineral processing papers deal with the utilization of coal fines. However, the research topics reflected in these papers are of less concern than the overall quality. Four and five years ago the number of learners entering the degree programmes represented by these papers showed a marked increase, which means that competition to have a paper published in this edition of the Journal is intense. Even if a particular topic is only of marginal interest to the reader, I would recommend that you read the abstract and glance through the paper to appreciate the high quality of the work of these young professionals.

Of course there is another and more disturbing side to the rising number of graduates. This comes at a time of great difficulty for the global mining industry, and nowhere more so than in Southern Africa. After decades of undersupply, we are now seeing graduates struggling to find employment. Even more surprising is that some of the best graduates, having received mining industry bursaries throughout their undergraduate studies, are being cut loose on graduation to find a job anywhere they can. The education they have received and the skills they have learnt generally equip them not only for their narrow specialization, but for a place in the workforce, often far from the mining industry. Once gone they are unlikely to return.

This raises the questions of whether the South African mining industry has a future, what sort of future will it be, and what sort of professionals will it require in order to be successful. This once great industry is at a low ebb as it struggles to come to terms with the forces imposed on it by the past decade or so, not least of which is the present slump in commodity

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prices. This was an industry, particularly the gold sector, that relied on brawn and where physical effort delivered the product. To remain competitive in the 21st century, mining and mineral processing has already changed significantly and will continue to do so. To achieve further gains in competitiveness, to continue the path to zero harm and to achieve the overall goals of sustainability, it will be brains and not brawn that will make the difference. It is all too easy to regard young graduates, at the start of their working lives, as a cost rather than an asset. While it is well understood that many graduates are unsuitable for a career within the narrow confines of ‘production’, the success of minerals industry companies is increasingly dependent on the ‘service departments’, where technical skills are in short supply.

Instead of employing young graduates only in production posts, while constantly bemoaning the low pass rates in industry certificates of competence, it could prove hugely beneficial for industry to broaden its vision of how to employ mining engineers and mineral processing graduates. Mine planning, rock engineering, mine ventilation, research and development, plant optimization, project management, and mechanization are some of the specialities that have a huge impact on the future of a company. While universities are seen by industry as a source of graduates with a broad-based education in their discipline, they also have a significant role to play in developing specialists through postgraduate qualifications. With a limited market for these courses some rationalization between universities is required, and this is where institutions such as the SAIMM and the other professional associations can help facilitate discussions.

We all know that mining is a cyclic business and also that ‘the darkest hour is just before dawn’. The high standard of papers in this edition of the Journal, coupled with the fact that these students have applied their minds to solving practical problems, gives a clear indication of the raw talent available to our industry.

Our industry must now have the foresight to develop this talent in order to ensure we are well positioned to take advantage of the next upturn in the commodities cycle.

H.R. Phillips

The Journal of The Southern African Institute of Mining and Metallurgy

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South African National Committee on Tunelling Young Members Group – ‘SANCOT – YMG’ oung professionals and the youth at large can make valuable contributions in the civil and mining industries. Like many other organizations within different industries, SANCOT has the responsibility of ensuring that there is effective youth involvement in all professional activities within the industry. This way, the youth are also able to make a meaningful contribution towards their professional and technical development.

At a meeting held on 14 January 2015 at AECOM offices in Centurion, the South African National Committee on Tunnelling approved the formation of a Young Members Group (’SANCOT –YMG’). Mr Lucky Nene was nominated and accepted as the chairman of this Young Members Group.

SANCOT–YMG has adopted its mandate from its mother bodies, SANCOT and the SAIMM, and is working very closely with the International Tunnelling Association (ITA) and the youth body of ITA, which is ITA-YM. This is to ensure alliance and compliance on various aspects that affect the young professionals. The mandate as adopted from the ITA-YM is structured as follows:

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To provide a technical networking platform within the ITA for young professionals and students To bridge the gap between generations and to network across all experience levels in the industry To create awareness of the tunnelling and underground space industry to new generations To provide young professionals and students with a voice in the ITA, including the Working Groups To look after the next generation of tunnelling professionals and to pass on the aims and ideals of the ITA.

Through general interactions via other professional platforms, young professionals have shown interest in this youth structure and a desire to take part. It is therefore envisaged that all interested companies would encourage their young professionals, within both mining and civil engineering, to have representatives within SANCOT-YMG. This participation and involvement is encouraged to extend beyond place(s) of work and will also include those young professionals that are at academic institutions.

The focus areas for SANCOT-YMG would be to mirror the mother body activities and objectives in a way that ensures fun, enthusiasm, and the ongoing participation of young professionals in all aspects of the mother body and industry at large. These focus areas are as adopted form the ITA-YM mandates and include the following: ­ Arranging events for international networking, and exchange of experience and technology between young professionals and students ­ Inspiring the young generation to join and actively participate in ITA ­ Encouraging member nations to establish domestic YM groups for each individual ITA member nation.

To date, SANCOT-YMG has embarked on a number of activities, including researching other existing professional youth organizations and groups in order to understand how they are structured, what their current involvement is, and where SANCOT-YMG can participate in the promotion of young professionals’ interests. To date these include ITA–YM, SAIMM–YPC, and CESA-YPF. As a way forward, SANCOT–YMG intends to embark on the following activities: a) Requesting assistance from the mother body in the formulation of the working committee/council b) Continue engaging with various young professionals’ organizations and other related stakeholders in an attempt to strengthen relationships and pursue youth interests c) Start implementation of the ITA-YM mandates in association with the SAIMM mandates d) Continue to participate in the activities of the ITA-YM, SAIMM-YPC, CESA–YPF, and other youth groups both locally and internationally.

All interested young professionals and those who would like to participate in general and offer assistance in the sustenance of this your professional entity are invited to contact the SANCOT–YMG chairman directly on Lucky.Nene@aecom.com or via Raymond van der Berg on raymond@saimm.co.za

L. Nene

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The Journal of The Southern African Institute of Mining and Metallurgy

closed off my March President’s Corner by making the point that we are in the ‘Age of Unicorns’ and that without strong math and science skills our future engineers and managers will not be adequately equipped to meet the needs and expectations of the national economy or for managing business complexity in the future. By way of example, a 2014 Department of Higher Education and Training report indicated that six of the top ten occupations in high demand were for graduated/certificated engineers (see the Table below, and of the remaining four occupations, three required higher technical training in engineering.

It will take significant efforts at all levels of our society to rectify ongoing imbalances in the Table I education system that seems ill equipped to Top 10 Occupations in high demand in South Africa motivate more learners to excel in maths and No. Occupational Title science and go on to study engineering at a senior level. However, it is all too easy to look 1 Electrical Engineer to the education system as having the sole 2 Civil Engineer responsibility to do this. Perhaps we should first 3 Mechanical Engineer look at how we, as parents, guide our children 4 Quantity Surveyor and nurture their inquisitive and analytical 5 Project Manager/Engineer minds from an early age. More should and can 6, 7 Finance Manager Physical and Engineering Science Technicians* be done to educate and inform parents about 8, 9 Industrial and Production Engineers* their educational responsibilities and change the Electrician culture which seems to imply that; once we have 10 Chemical Engineer packed the kids off to school, that is the end of our educational responsibility as a parent. t This edition of the Journal showcases some of our industry’s young engineering talent, who have risen to the above challenge, and publishes the work they are doing to better understand some of the technical issues facing both our mining and extractive sectors. Most of the authors either have started, or are about to start, their working careers. The SAIMM has two main initiatives through which it strives to contribute towards the ongoing development of engineers, specifically for the mining industry: 1. The SAIMM Scholarship Trust Fund. This Trust channels financial assistance to underprivileged and talented undergraduates. We already have numerous case studies of lives being transformed. The SAIMM desperately needs the support of its member’s contributions to this cause, simply because the results are so immediate and measureable 2. The Young Professionals Council: Many of you reading this article will recall with mixed memories and emotions your first two to three years working for your first boss who was both task-driven and not a particularly good coach … There are so many young engineers that do not handle this transitional period well, and our Young Professionals Council has been tasked specifically to find ways of staying close to graduates and diplomates at this critical start to their careers, in order to offer friendly advice, support, and guidance.

Technology continues to drive the demand for engineers. Safety and efficiency are drivers of mechanization and automation in mines and manufacturing around the world; smartphones, cheap sensors, and cloud computing have enabled a raft of new internet-connected services that are infiltrating the most tech-averse industries—Uber is roiling the taxi industry; Airbnb is disrupting hotels. Perhaps ongoing research towards continuous mining systems will also reinvent the mining industry? Certainly, technology entrepreneurs are exploiting the new technology opportunities.

So what about the ‘Age of Unicorns’?

This also relates to the pace of technology change that is driving the demand for engineers of all disciplines. You will recall that unicorns are mythical creatures that existed in people’s imaginations? Well, the billion-dollar tech start-up was supposed to be the stuff of myth; neither Google not Amazon were in the billion dollar league on start-up (Aileen Lee coined the term unicorn as a label for such corporate creatures.). There are now (according to Fortune) more than 80 companies with more than this value at start-up. And it is accelerating: in 2013 there was one company with a start-up valuation of $10 billion but today in 2015 there are eight (including Uber, the on-demand car service worth $41.2 billion.

Its valuation is higher than the market capitalization of at least 70% of the companies in the Fortune 500!)!

Why does this all matter? Because these start-ups are not in the ‘classical’ engineering space but are pulling the top math and science talent from the built environment. We need to look after our own and ensure that even when the mining industry is in challenging times we do not neglect to invest in our young professional engineers.

t’s iden s e r P er Corn

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J.L. Porter President, SAIMM

Handover of model stope to Wits School of Mining Engineering 09 March 2015 – Johannesburg: A life-size mining stope panel was handed over by New Concept Mining (NCM) to the Wits School of Mining Engineering on 6 March 2015, to help students learn about stoping activities through a better visualization of how a real mine looks.

The stope panel – sponsored to the tune of R250 000 by NCM – is part of a range of simulated facilities sponsored and developed at the School’s premises on West Campus, in partnership with companies active in the mining sector such as Aveng, Gold Fields, and Sibanye. These include a mine tunnel, mine shaft steel work, and a lamp room.

Professor Cuthbert Musingwini, newly appointed Head of the School of Mining Engineering at Wits, said: ‘We are delighted to add this new facility to our School’s resources and grateful to be partnering with far-sighted stakeholders like NCM who share our dedication to skills and technology development.’

NCM marketing director Brendan Crompton said the sponsorship of the model stope panel was driven by NCM’s commitment to safety, efficiency, and productivity in South African mines. The SA-based company is a market leader in narr0w-reef stope support products, and has expanded into a number of countries worldwide.

‘As a quality-focused company rooted in South Africa, we recognize that the future of our mining sector is built on the calibre and skills of graduates from institutions like Wits University,’ said Crompton. ‘Partnering with the School of Mining Engineering at Wits is one of the ways that we contribute to sustainability and safety in mining, especially as we both prioritize technological innovation as a key factor in the success of the sector.’

Measuring some seven metres in length, the model stope was constructed from a metal Professor Cuthbert Musingwini, Head of framework, mesh, and concrete. Sculptor Russell Scott used various materials and techniques, the School of Mining Engineering, Wits, including hand-packed cement and layers of paint, to achieve the realistic effect of a working opening the official handover stope face in an underground platinum mine.

The panel dips at 10 degrees, has a stoping width of one metre, and extends some three metres on strike. It has been equipped with various items of support infrastructure to demonstrate to students the variety of technologies employed underground. These include timber props, timber packs, rockbolts, and safety nets suspended near the working face.

NCM has contributed roof support equipment both from its own range of products and from other sources. It is also making available some of its electronic monitoring and warning devices in the stope, augmenting the School’s focus on digital remote monitoring technologies to enhance safety on mines.

Like the recently completed model mine tunnel, the stope panel is situated in the basement of the School of Mines premises, where it incorporates one of the building’s beams as a geological feature.

Professor Fred Cawood, former Head of the School, initiated construction of the stope panel as part of his digital mine research at Wits Mining. These simulated facilities form part of the ‘digital mine’ environment which is providing invaluable tools for learning and research, bringing a real mine experience to mining engineering students at Wits.

‘Most of the 200 first-year students we welcome each year are straight from school and have never been in a mine before,’ said newly appointed Head of School Professor Cuthbert Musingwini. ‘Although mine visits are arranged from time to time, this facility gives easy access to students – so that they can visualize and test what they are studying theoretically.’

‘While the facility is invaluable for our teaching work, it will also be made available to our research students as they push the boundaries of productivity with digital and other technology in mining,’ said Professor View looking up-dip showing the edge of the gully, Cawood. ‘Now more than ever, South Africa needs to encourage and facilitate mechanical props, and gully pack research that can stimulate our mining sector; through facilities like these, Wits School of Mines is showing its commitment to doing that.’

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New head of Wits mining school announced 9 March 2015 – Johannesburg: Professor Cuthbert Musingwini has been appointed head of the University of the Witwatersrand’s School of Mining Engineering.

Having lectured at Wits since 2004, Professor Musingwini has over 20 yeaars of experience in the mining sector – including mine production management and planning, consulting, and academia. He began his career in the Zimbabwean gold mining industry as a research fellow – and later a lecturer – at the University of Zimbabwe.

He is Senior Vice-President and Honorary Treasurer of the Southern African Institute of Mining and Metallurgy (SAIMM), a Fellow of the SAIMM, a registered professional mining engineer with the Engineering Council of South Africa (ECSA), and holds a PhD in Mining Engineering from Wits. He is a Managing Editor of the International Journal of Mining, Reclamation and Environment published by the Taylor and Francis Group (UK). He was awarded a National Research Foundation (NRF) C3 rating in 2014, and has published and presented extensively both locally and abroad.

S. Braham

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Wits-SRK link boosts rock engineering skills 17 March 2015 – Illovo, Johannesburg: Collaboration between Wits University’s School of Mining Engineering and consulting engineers SRK Consulting is nurturing scarce rock engineering expertise, benefiting the mining and other sectors in Africa and beyond.

For the past decade, SRK has partnered with Wits through providing financial support for selected students in the Wits School of Mining Engineering’s postgraduate rock engineering research programme, as well as internship opportunities within the firm. ’The bursary programme has allowed some of our top students to specialize in rock engineering, which is a key discipline for mining but which for various reasons attracts relatively little postgraduate interest among graduates,’ said Professor Cuthbert Musingwini, Head of the School.

The scheme was initiated by Professor Dick Stacey, then Centennial Professor of Standing: Philani Mpunzi Rock Engineer SRK Consulting (SA), Prince Mulenga Student Intern SRK Rock Engineering at Wits, who approached a number of companies and organizations Consulting (SA), William Joughin Principal in the mining sector to seek their help in dealing with numerous requests from bright Geotechnical Engineer and Partner SRK Consulting (SA), Prof Emeritus Dick Stacey of the Wits School of but under-resourced students wanting to undertake MSc studies.

Mining Engineering ‘I was delighted when SRK took up this challenge, and also offered to take in some Seated: Joseph Mbenza Muaka, Rock Engineer SRK Consulting (SA) and Prof Cuthbert Musingwini, Head of the students as interns,’ said Professor Stacey, who spent 25 years of his career of the Wits School of Mining Engineering working at SRK and is today Professor Emeritus at the Wits School of Mining Engineering. ‘Internships are hugely valuable for postgraduate students, giving them real-life work experience and practical mentoring while leaving space for them to complete their studies.’

SRK partner and principal consultant William Joughin has been integrally involved with the bursary students who have internships at the firm.

‘This partnership helps us identify the best MSc students to assist us with many of our projects,’ said Joughin, ‘and it is heartening to see how they develop their skills during their time with us.’

He said that ten students had gone through the bursary-intern route at SRK, while another six SRK employees have completed – or are busy with – an MSc at the Wits School of Mining Engineering. Two more have been employed after they earned their MSc degrees.

Highlighting the quality of the candidates who have benefitted from the scheme, Joughin said: ‘Three of the previous students have received medals from the Southern African Institute of Mining and Metallurgy (SAIMM) for papers based on their MSc project work carried out at SRK. Some have also received awards from the South African National Institute of Rock Engineering (SANIRE) for top marks in the Chamber of Mines Rock Mechanics certificate.’

Mining engineer Joseph Mbenza Muaka worked on mining operations and at Mintek before completing his MSc at Wits – with support from another bursary provider, Coaltech – while working as an intern at SRK. He recently presented a paper at the Southern Hemisphere International Rock Mechanics Symposium (SHIRMS) on numerical modelling.

SRK’s status as a leading consulting firm with strong roots in technical excellence also allows interns to be exposed to cuttingedge investigations. Intern Prince Mulenga, currently busy with his MSc at Wits, will be involved in a project funded by the Safety in Mines Research Advisory Committee (SIMRAC) at the Mine Health and Safety Council – also an important partner of the Wits School of Mining Engineering.

The internship system has allowed some of the MSc graduates to progress within SRK and to become mentors to the newer interns. Philani Mpunzi, who completed his studies in 2011 under Professor Stacey, is now a specialist 3D modeller for SRK and helps interns to make the most of their time while optimizing their contribution. Mpunzi and another SRK/Wits student, Tazibana Moyo, co-authored a paper on their MSc research at the SHIRMS conference.

‘Having worked in Zimbabwe’s mining sector for six years as a production supervisor and mine planning engineer, I appreciate being able to share my experience while contributing to the development of young rock engineers,’ said Mpunzi.

While SRK does not have capacity to absorb all its interns, there is considerable opportunity to progress through the ranks. One of the first interns, Robert Armstrong, joined SRK as a research student in 2001 and a full-time engineer in 2005; last year, he was promoted to associate partner.

‘Perhaps one of the most valuable aspects of studying and doing research while engaged by SRK is the ability to get relatively easy access to highly experienced experts in fields like rock engineering,’ said Armstrong.

Another positive element of the Wits-SRK partnership is the role played by SRK’s rock engineering experts in the postgraduate courses themselves, according to Professor Stacey.

‘At least seven of SRK’s best technical minds have contributed to our MSc courses as guest lecturers,’ he said. ‘There has also been considerable time invested by SRK experts Peter Terbrugge and Robert Armstrong geologist and William Joughin as external examiners for these courses.’ associate partner SRK Consulting (SA) Professor Musingwini said the partnership indicates the way forward for the mining sector in South Africa and beyond our borders; it has already contributed well-qualified rock engineers to companies outside SRK and to countries across Africa and abroad. ’It is vital that academia, industry, and the public sector work closer together if we are to successfully overcome the skills challenge that mining faces and invite other consulting companies to partner with us in other areas of specialization on models similar to the Wits-SRK link,’ he said.

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Student Papers Re-aligning the cutting sequence with general support work and drafting a support sequence at Simunye Shaft

by K. Lombard . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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Roof support awaiting time (RSAT) is the potential production time lost while waiting for roof support to catch up with the continuous miner. The causes of excessive RSAT were investigated, and a number of solutions were identified, of which the use of hard roof drill bits as standard was shown to be the most effective.

Mining through areas affected by abnormal stress conditions at Syferfontein Colliery

by C. Legote . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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This paper investigates the conditions leading to the indefinite termination of production in four critical primary panels at an underground coal mining operation. The observed shortcomings in the mining approach were identified, and a strategy is proposed for mining through the affected panels.

A critical evaluation of the water reticulation system at Vlaklaagte Shaft, Goedehoop Colliery

by R. Lombard . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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The various factors that contributed to the high water-related downtime, which seriously affected production, were investigated. The water reticulation system was reviewed, and the current and future underground pipe layout and water requirements were determined for the shaft.

Optimization of shuttle car utilization at an underground coal mine

by P.R. Segopolo . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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The purpose of the project was to convert the current shuttle car utilization on an underground coal mine to best practices by focusing on change-out points and tramming routes, which have a major influence on shuttle car away times. Shuttle car utilization can be improved by balancing the number of cars with the number of open splits and the mining sequence.

Explosives utilization at a Witwatersrand gold mine

by M. Gaula. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

This investigation examined the properties of explosives, mine standards and recommendations for usage along with the historic relationship between the quantities of explosives used and the production output. This was then compared to the expected quantity of explosives required per unit of production. The biggest contributor to the apparent under-utilization of explosives was found to be the limitations of the system that tracks the usage of explosives underground.

These papers will be available on the SAIMM website

http://www.saimm.co.za

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Critical investigation into the problems surrounding pillar holing operations

by J.P. Labuschagne, H. Yilmaz, and L. Mpolokeng . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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An investigation into pillar cutting practices was carried out at a platinum mine in order to improve the compliance for pillar cutting. The findings suggest that the pillar strength problem lies with the implementation of the design rather than the pillar design itself.

LHD optimization at an underground chromite mine

by W. Mbhalati . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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The factors preventing the load haul dump machines (LHDs) from tramming the target tonnages at an underground chromite mine were investigated. Simulations showed that production improvements of more than 100% could be obtained by reducing the one-way tramming distances and optimizing LHD utilization.

The viability of using the Witwatersrand gold mine tailings for brickmaking

by M. Malatse and S. Ndlovu . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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This work examines the use of gold mine tailings, in various ratios with cement and water, for brickmaking.

The bricks were tested for unconfined compressive strength, water absorption, and weight loss. The results indicated that gold mine tailings have a high potential to substitute for the natural materials currently used in brickmaking.

Evaluation of some optimum moisture and binder conditions for coal fines briquetting

by P. Venter and N. Naude . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

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The optimum binder and moisture additions to produce a mechanically strong briquette from coal fines were investigated using two different binders.

Air drying of fine coal in a fluidized bed

by M. Le Roux, Q.P. Campbell, M.J. van Rensburg, E.S. Peters, and C. Stiglingh. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .

The rate of moisture removal from coal particles in a fluidized bed under a range of operating conditions was investigated. It was found that the relative humidity of the drying air has a greater effect than temperature on the drying rate, even at temperatures as low as 25째C. The energy efficiency of the fluidized bed compared favourably with other thermal drying methods.

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/v115n4a1 ISSN:2411-9717/2015/v115/n4/a1

Re-aligning the cutting sequence with general support work and drafting a support sequence at Simunye Shaft

by K. Lombard* The work presented in this paper was carried out as partial fullfilment for the degree BEng (Mining Engineering)

‘Roof support awaiting time’ (RSAT) is a term used at Goedehoop Colliery’s Simunye Shaft to describe the potential production time lost due to the continuous miner (CM) standing idle waiting for roof support to catch up.

Investigations revealed that in 2013, Simunye Shaft had approximately 1400 hours of RSAT, which suggests that the mine could have potentially produced an additional 280 000 t of coal. This project consisted of two parts. Firstly, the causes of the high RSAT and means to improve the situation were investigated. Secondly, as insisted by mine management, the CM cutting sequence was investigated as a possible cause of high RSAT. Machine-related challenges due to the roofbolter installing support too slowly, geological conditions (mostly hard roof conditions and slips), logistical challenges pertaining to the CM cutting sequence, man-related challenges related to operator fatigue, re-support, operator inexperience, and the absence of support targets were identified as main contributors to RSAT. Furthermore, results showed that the roofbolters in the sections at Simunye Shaft are slower than the CMs. A target of 28% reduction in RSAT was set. Experts from Kennametal and Fletcher were consulted to find solutions for the identified causes. In total, eight solutions for RSAT were identified, but the solution that contributed most significantly to reducing RSAT was to use hard roof drill bits as a standard product at Simunye Shaft. Calculations showed that by using hard roof drill bits, RSAT can be reduced by 43%, which is more than the specified 28% target.

The cutting sequences at Kriel, Greenside, and Simunye Shaft, together with three newly developed cutting sequences, were simulated using the UCMS (Underground Coal Mining Simulation) program. A re-aligning principle was incorporated into the newly developed cutting sequences to align the cutting sequences to general support work and to reduce RSAT. A decision matrix revealed that a cutting sequence in which boxing takes place in R3 (third road to the right of the belt road) and in which the realigning principle has been incorporated will be the best option for Simunye Shaft. The recommended cutting sequence will lead to a 5% increase in production.

Keywords roof support awaiting time, CM cutting sequence, simulation, hard roof drill bits, support sequence.

Mine background and general information Simunye Shaft is an Anglo Thermal Coal underground mining operation. The mine is situated approximately 41 km from Emalahleni, Mpumalanga Province. Simunye Shaft has five bord and pillar sections, four of which are continuous miner (CM)-shuttle car sections and one is a CM-FCT (flexible conveyor train) section. The No. 4 Seam is mined at an average seam height of approximately 3 m.

At Simunye Shaft, the floor, comprising largely sandstone or a sandstone/siltstone The Journal of The Southern African Institute of Mining and Metallurgy

Project background ‘Roof support awaiting time’ (RSAT) is a term used by Goedehoop Colliery to describe the potential production time lost due to the CM standing idle waiting for roof support to catch up. There are two types of RSAT, namely operational RSAT and engineering RSAT. Operational RSAT (Figure 1) is driven by operational processes e.g. the roofbolter falling behind the CM due to adverse geological conditions, damaged roofbolts, material shortage, cutting out of sequence (which may lead to logistical problems that will prevent the roofbolter from finishing support in time), etc.

Engineering RSAT is potential production time lost due to roofbolter breakdowns. Support awaiting time has proven to be a major bottleneck in production at Simunye Shaft. As illustrated in Figure 1, support awaiting time amounted to 1700 and 1400 hours in 2012 and 2013 respectively. This means an additional 280 000 t could potentially have been produced in 2013. On average, almost 14% of available in-section production time was lost due to operational RSAT. The mine lost a potential R125 million in revenue, and although this was less than in 2012, it was still an enormous loss. The RSAT of the No. 4 Seam CM-shuttle car sections amounted to approximately 72% of the total RSAT, and therefore this project covers only No. 4 Seam CM-shuttle car sections at Simunye Shaft.

* Faculty of Engineering, Built Environment and Information Technology, Department of Mining Engineering, University of Pretoria. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Jan. 2015 VOLUME 115

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Synopsis

combination, is expected to be reasonably competent. The immediate roof, however, consists of an interlaminated unit of shale and siltstone and may present roof stability challenges. The roof encountered during mining varies from soft to hard, and hard roof conditions may result in support challenges (Mathetsa, 2013).

Re-aligning the cutting sequence with general support work

Figure 1 – Total downtime hours awaiting roof support – Simunye Shaft

Figure 2 – Underground logbook RSAT data analysis 2013

Support is a major component off the production process.

If three faces are left unsupported (according to Anglo American Thermal Coal standards) the CM has to wait for support before production can commence. Therefore, to improve section productivity, it is vital to make the timeous installation of roofbolts a priority.

The question arises as to what may be the causes of high RSAT. It was the author’s goal to identify the main causes of the high RSAT at Simunye Shaft and to suggest strategies that will make the timeous installation of roofbolts a priority.

Based on underground logbook data, the causes of RSAT can be divided into five main categories (Figure 2): ® Problems related to the roofbolter being too slow or on breakdown (machine problems) ® Geological conditions: hard roof conditions, conditions in which slips are encountered and in which oslo straps have to be installed ® Logistical challenges (when the roofbolter is blocking the CM) due to the cutting sequence not taking the interaction between the CM and roofbolter into account ® Man-related challenges – operators supporting too slowly or arriving late for work. Re-support and material shortage also fall into this category. Operators need to re-support when roofbolts are damaged during the roofbolting process or when the spacing between the roofbolts is inadequate ® Infrequent events, includes when support is updated, a temporary support jack has to be installed, and when the roofbolter has to wait for the LHD to complete sweeping.

As indicated in Figure 2, machine, geology, and logistical challenges are the main contributors to RSAT, contributing 82% of the problem.

It should be noted that the logbook data was very incomplete and that more than half of the RSAT could not be accounted for. Only 190 data points out of total of 800 were used as a result. Owing to the incompleteness of the data, a survey was conducted among underground workers to obtain a better understanding of the causes of the high RSAT. Twelve surveys were completed (results depicted in Figure 3).

The installation of oslo straps and hard roof conditions were identified as main causes of the high RSAT. Most of the shift bosses raised the concern of a lot of new inexperienced workers in their sections, and one shift boss mentioned that 60% of his operators were new and had not received

sufficient ff on-the-job training from f the retiring workers. Most of the workers mentioned that the CM is faster than the roofbolter and that this is the cause for the high RSAT. Engineering breakdowns were also mentioned as a problem, and may be attributed to the fact that maintenance on roofbolters is not seen as a big priority.

It is important to identify and investigate the main aspects that contribute to RSAT so that RSAT can be reduced. By taking the analysis in Figure 2 and the survey results in Figure 3 into consideration, the following will be investigated: ® Machine considerations – means by which to increase the speed of roofbolt installation. Slow roofbolting is the main contributor to RSAT ® Geological conditions (including hard roof conditions and slips) – this is the second highest contributor to RSAT ® Logistical issues – means by which to improve the CM cutting sequence ® Man-related challenges – operator fatigue as a contributor to RSAT ® Other challenges – re-support ® Operator inexperience ® The implementation of support targets (bolts installed per shift).

Engineering breakdowns will not be investigated, because these are machine-related and can be prevented only by means of a better maintenance plan (increase in maintenance time, maintenance staff, and more reliable equipment).

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Figure 3: Survey results: causes of RSAT The Journal of The Southern African Institute of Mining and Metallurgy

Re-aligning the cutting sequence with general support work Table I

Identified objectives and methodology to attain the objectives Objective

Methodology

Conduct a root cause analysis of the high RSAT at Simunye Shaft

• • •

Reduce RSAT by 28%

• • •

Re-align the CM cutting sequence to general support and formulate a support sequence

•

• • •

• Support targets

• • •

Investigating the engineering breakdowns off the roofbolter f as a contributor to RSAT’s is, however, suggested as a topic for further work. It should also be noted that purchasing new roofbolters (so that there are two roofbolters available per section) was eliminated as a solution, seeing that Simunye Shaft does not have the capital to purchase new roofbolters.

The high RSAT called into question the current CM cutting sequence, which was developed with a main focus on the CM. Will a new cutting sequence with a support approach improve RSAT? A deeper investigation into the cutting sequence currently employed at Simunye Shaft will be performed as the mine sees it as a priority to optimize its cutting sequence by re-aligning it to general support work and thereby reducing RSAT.

A survey was undertaken among underground workers to obtain their opinions regarding the causes of the high RSAT Underground logbook data was analysed to determine the main causes of the high RSAT The underground logbook data analysis and the survey results enabled the main areas for improvement to be identified.

It was necessary to determine the time it takes to support a 9 m heading in order to determine what effect new technology may have on RSAT Experts from Fletcher and Kennametal were consulted regarding new technology that could be implemented to reduce RSAT The cost of the various solutions and initiatives to reduce RSAT was taken into consideration to ultimately make recommendations.

Underground observations and interviews with underground workers assisted in identifying the main areas of concern regarding to the current cutting sequence at Simunye Shaft Experts from other mines were consulted to obtain information with regard to their cutting sequences Rock mechanical and ventilation standards were taken into consideration when cutting sequences were developed A UCMS simulation program was used to simulate the different cutting sequences (three proposed cutting sequences and cutting sequences of other mines) in order to select the best cutting sequence for Simunye A logical analysis of the optimal developed CM cutting sequence assisted in drafting a support sequence.

Support targets were set up for various underground scenarios (when a hard roof or slips are encountered and for normal conditions) The advance rates of the roofbolters were determined through time studies and industry data The effective production time per shift (approximately 3 hours per shift) and the support advance rate for the various scenarios were used to determine support targets.

The literature study consisted off five f parts. Firstly, a standard support sequence was described. This assisted in understanding the roof support process and provided a starting point from which improvements could be made.

Secondly, the changes brought about in the support process when slips are encountered were described. Slips are one of the causes of RSAT – support spacing is reduced when a slip is encountered, which increases the time required for installing support. Thirdly, the main causes of RSAT at Kriel Colliery were investigated to strengthen the motivation for the study. Fourthly, solutions that may reduce RSAT were investigated. Finally, the cutting sequences employed at Simunye Shaft, Greenside Colliery and Kriel were described to illustrate where improvements in Simunye Shaft’s cutting sequence could be made.

Objectives and methodology The objectives that were identified during the course of the project as well as the methodology used to meet the objectives are set out in Table I.

Literature survey

The Journal of The Southern African Institute of Mining and Metallurgy

Figure 4 – Plan view of the ideal roofbolt installation sequence for a double-boom roofbolter (Van Staden, 2014) VOLUME 115

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An extensive investigation into work that has already been done to reduce RSAT at underground coal mines, yielded only limited information. This may be attributed to the fact that, in general, underground data – for example the number of roofbolts installed and amount of drill steels used per shift – is not recorded in an organized or accurate manner.

Therefore, the causes of the RSAT cannot be pinpointed easily. The analysis of underground logbook data in order to determine the causes of RSAT is a time-consuming process.

Re-aligning the cutting sequence with general support work

Figure 5 – Plan view of the change of roofbolt spacing from normal conditions to when a slip is encountered

Figure 7 – Oslo strap holder (Steyn, 2013)

Figure 6 – RSAT breakdown at Kriel Colliery for 2012 (derived from underground logbook data)

Standard roofbolt installation sequence In Figure 4, the red dots indicate where roofbolts are installed, and the numbers indicate the sequence in which the roofbolts are installed (mining takes place from left to right in the figure). Looking in the direction of mining, support starts at the beginning of the heading at the far left-hand and far right-hand side simultaneously (two roofbolts are installed at the same time) and then proceeds to the inner left-hand and right-hand side. The same installation sequence is employed at Simunye Shaft.

Slips If slips are encountered at Simunye Shaft, the spacing between two consecutive lines of support is reduced from 1.5 m (normal conditions) to 1 m as illustrated in Figure 5. It should be noted that, in the figure, mining takes place from the bottom upward. If multiple slips are encountered, oslo straps need to be installed in addition to the reduced spacing.

RSAT at Kriel Colliery By analysing the RSAT at another colliery, the significance of the problem can be emphasized and the motivation for the study strengthened. Kriel Colliery had 1401 hours (almost exactly that of Simunye Shaft) of operational RSAT in 2012.

This illustrates that other collieries have the same types of problems and a mind-shift is needed to overcome the problem – roof support needs to become a higher priority. Figure 6 illustrates the breakdown of the RSAT at Kriel Colliery for 2012. Machine-related challenges are documented as the

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highest contributor to RSAT, mainly due to the roofbolter being too slow. Logistical issues related to the CM cutting sequence are 10% lower than that of Simunye Shaft. It should, however, be noted that almost 64% of the data points in this analysis could not be used as they were recorded as ‘n/a’, or unaccounted for. It is clear that RSAT is not only a problem at Simunye Shaft, but also at the other collieries in South Africa. It is of great importance to solve this problem to ultimately enable collieries to increase their efficiencies.

Possible solutions The oslo strap holder According to Steyn (2013), a consultant at Fletcher, oslo strap holders (which are attached to the roofbolters at Greenside Colliery) have the potential to reduce the time to install an oslo strap by more than one minute. The oslo strap holder (Figure 7) assists the roofbolter operator to position the oslo strap (which normally takes a considerable amount of time) and in doing so increases safety significantly.

Standardizing the drill bits used at Simunye Shaft to hard roof drill bits A hard roof decreases the penetration rate and bit life (bits will have to be changed more frequently). Both the reduction in penetration rate and the excessive replacement of drill bits lead to an increase in support time. Hard roof drill bits are available and are used by the mine for hard roof conditions.

The hard roof drill bits should increase the life of a single drill bit and may increase the penetration rate.

Table II contains information with regards to the penetration rates and costs of the different drill bits that can be supplied to Simunye Shaft and the different roof conditions encountered. The figures are estimated and may The Journal of The Southern African Institute of Mining and Metallurgy

Re-aligning the cutting sequence with general support work Table II

Relationship between type of drill bit used and the installation time of a roofbolt for various roof conditions (Bosch, 2013)

KCV4 1 RRWT SV119AE K30 12EX02 PROBORE1 HSVSL

Cost (R)

Installation time of Installation time of one roofbolt in one roofbolt in standard roof hard roof conditions (min) conditions (min)

Life of drill bit penetrating coal (number of holes drilled before replacement)

Life of drill bit penetrating sandstone (depends on hardness) (number of holes drilled before replacement)

Life of drill bit penetrating coal with intrusions (depends on hardness and the intrusion) (number of holes drilled before replacement)

Life of drill bit penetrating quartzsite or harder roof (number of holes geometry of drilled before replacement)

37.20 43.71

2.5 1.6-2.5

8 6-8

200 100

0-14 5-14

3-7 2-5

1-4 1-5

60.16

1.6-2.5

6-8

300

25

0-10

0-7

vary with operator skill, geology, and the consistency of adjusting the roofbolter’s settings if roof conditions change.

The KCV4 1 RRWT (Table II) is the current drill bit employed at Simunye Shaft for normal roof conditions. If a hard roof is encountered, the SV119AE K3012EX02 drill bit (Table II) is used. The PROBORE1 HSVSL (not used at Simunye Shaft) has very good heat resistant properties, which increases drill bit life, but its high price makes it unsuitable for use as a standard product (Bosch, 2013).

When used in hard roof conditions, the KCV4 1 RRWT drill bits (which are designed for normal roof conditions) and drill steel heat up rapidly and melt into the adapter. Removing the drill steel from the adapter can easily take 30–60 seconds (Bosch, 2013). By standardizing to hard roof drill bits, the installation time per roofbolt can be decreased by approximately 30 seconds (Table II). It should be noted that if KCV4 1 RRWT drill bits are used for hard roof conditions, the time to install a roofbolt may increase to 8 minutes (Table II).

The torque indicating system According to Sinden (2013), the installation quality of a support system is directly related to the performance of the machinery used to install the bolts. Statistics shows that only 20% of all bolters have torques set within the correct

Figure 8 – Current cutting sequence at Simunye Shaft. Green blocks (numbers 1– 28) are the first sequence and the repetition of the sequence is indicated by the red blocks (numbers 29– 56). B/R – belt road, FB – feeder breaker, L1 – first left road, R1 – first right road etc.

Triangular shapes in R3 indicate where boxing takes place The Journal of The Southern African Institute of Mining and Metallurgy

Table III

Average time (actual stopwatch time) to support a 9 m heading (Van der Merwe, 2012) Activity

Total average (h:min:s)

TRS up (re-position time ) Total Drill Insert extension drill rod Change to bolting equipment Bolting Re-position for next bolt TRS down (time per row ) Bolts per heading Time per heading Relocation time

00:00:36 00:00:50 00:00:21 00:00:49 00:00:17 00:00:50 00:07:58 25 (9 m) 00:50:46 00:20:25

Average LH (h:min:s)

Average RH (h:min:s)

00:00:48 00:00:14 00:00:50 00:00:16 00:00:47

00:00:53 00:00:29 00:00:47 00:00:18 00:00:53

Newton-metre range (200–350Nm). Sinden also mentions that the operator torques the roofbolt according to what he thinks or sees is right, and that that is the reason why common faults such as over-torque (flattened washers) and under-torque (loose washers) conditions occur. In the event of over-torque, the bond between the resin and bolt may be damaged, the washers may be deformed, the nuts may be rounded, and roofbolt threads may be damaged. Undertorque results in loose washers, incorrect mixing of resin, not breaking the shearing pin, and not flattening the torque indicator (Sinden, 2013).

The torque indicating system was designed to avoid substandard torque (which may require re-support of roof bolts). The torque indicating light is clearly visible during operation and ensures accurate bolting torque. The operator, miner, and technician can also see when the torque of the roofbolter drill chuck is not optimal (Sinden, 2013).

The torque indicating system has a data logging feature that can capture data such as the torque, time, date, the number of roofbolts installed, spinning time of resin, and holding time before a roofbolt reaches torque (to help estimate accurate roofbolt and resin usage). The torque indicating system will therefore result in the better management of roofbolting crews, as managers will know how the crew has performed (number of bolts installed per shift). The system will also reduce RSAT resulting from reVOLUME 115

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Drill bit type

Re-aligning the cutting sequence with general support work support being required. The torque indicating system is a new product on the market and its advantages can be summarized as follows (Sinden, 2013): ® The torque indicating light is highly visible ® If the system fails to operate correctly, the operator can identify and report immediately ® Accurate torque on every roofbolt is ensured ® The system helps to overcome human errors and faulty hydraulic systems ® The system is cost-effective ® It reduces machine working hours ® Time on re-installation is saved ® Labour and re-installation costs are reduced ® The system eliminates the re-occurrence of over- and under-torque.

The MCS roofbolter monitoring system MCS offers two options for data recovery (operating time, tramming time, downtime, and number of roofbolts installed) with the roofbolter system. The first system, which is currently in use at Anglo Thermal Coal sites, is a flash card system. The operator is responsible for inserting the flash card at the beginning of the shift and returning it to the control room for processing at the end of shift. The second is a Wi-Fi system, where the onboard data collection unit communicates to the node which is integrated into the mine’s communication backbone, allowing file transfer to the control room. This can be done in two ways – firstly, by means of two nodes, and secondly by means four nodes. The advantage of installing more nodes is increased coverage.

The MCS system and torque indicating system are similar.

The biggest difference between the two systems is cost – the torque indicating system is more cost-efficient.

The auto-bolter According to Steyn (2013) a roofbolter operator handles approximately 1.5 t of steel per shift and makes approximately 14 lever movements per roofbolt installed. The operation of a roofbolter involves strenuous tasks, and the time to install a roofbolt can increase from 2.5 minutes at the beginning of a shift to approximately 10 minutes at the end of a shift as a result of operator fatigue. The weighted average time to install a bolt is 6.75 minutes. This was calculated by increasing the time to install a bolt linearly every 30 minutes over a 3 hour (effective operating time) period.

The autobolter technology, which is currently being implemented at Greenside Colliery, may eliminate this problem. The autobolter has the following advantages (Steyn, 2010): Safety: ® Reduces the number of accidents related to the handling of roofbolts and the operation of the roofbolter ® Moves the operator to a safer position.

Productivity: ® Reduces operator fatigue ® Ensures the consistent installation of bolts.

Reliability: ® Removes the human factor ® Records roof mapping information ® All bolts are installed to the same standards and procedures.

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Engineering: ® The autobolter has various pressure settings for feed and rotation ® It is difficult to tamper with the machine setup ® Pressures and sensors are displayed on a display screen.

The autobolter ensures that each bolt is installed to the correct standard (re-support is eliminated) and operator fatigue is eliminated completely as the roofbolter is remotecontrolled and the operators do not have to handle the heavy roofbolts. However, the cost of the autobolter – approximately R14 million –eliminates it as a solution as the mine’s budget does not cater for the purchase of new roofbolters.

Current cutting sequence employed at Simunye Shaft It is necessary to analyse and evaluate the effectiveness of the cutting sequence currently employed at Simunye Shaft, as it was developed with a main focus on the CM (a support sequence was never developed). A support sequence can be described as the sequence in which the roofbolter supports the headings and splits cut by the CM. The current support sequence employed at Simunye consists of the roofbolter following the CM sequentially as far as possible. The shortcomings of the cutting sequence, with regard to support have to be identified in order to identify solutions to the problem of RSAT. The current cutting sequence employed at Simunye Shaft is illustrated in Figure 8. The green blocks (numbers 1–28) are the first sequence and the repetition of the sequence is indicated by the red blocks (numbers 29–56).

The belt road (B/R), the first left road (L1) to the fourth left road (L4) and the first right road (R1) to the fourth right road (R4) are indicated. The triangular shapes in the R3 indicate where boxing takes place. Boxing is when a triangular shape is cut into the coal face to make it easier for the CM to manoeuvre when it is cutting cuts number 4 and 1 of each sequence. The feeder breaker (FB) is also indicated in the figure. Simunye Shaft’s cutting sequence ends in R2, which is close to where the following cutting sequence starts. Boxing takes place in R3 so that through ventilation is established as quickly as possible.

Cutting sequences at other collieries By investigating different cutting sequences employed by other mines (with more or less the same pillar sizes as Simunye Shaft), an optimal cutting sequence for Simunye Shaft can be developed.

Greenside Figure 9 illustrates the cutting sequence employed at Greenside Colliery. Boxing takes place in R1 and the cutting sequence ends at the far left-hand side.

Kriel Figure 10 illustrates the cutting sequence followed at Kriel.

Boxing takes place in R1 and the cutting sequence ends close to where the following cutting sequence starts.

In summary the following concepts can be incorporated into a cutting sequence: ® To box in R1 (reduce cable handling efforts and time) ® To box in R3 (establish through ventilation as soon as possible) ® To end the cutting sequence at the far left-hand side The Journal of The Southern African Institute of Mining and Metallurgy

Re-aligning the cutting sequence with general support work

Figure 9 – Cutting sequence at Greenside Colliery (Odendaal, 2014)

Figure 11 – Roofbolter exposed to water from the scrubber fan of the CM and dust

ffaces are left f unsupported, the CM has to wait ffor support to catch up. The most important ventilation standard that needs to be taken into consideration is that an air speed of 1 m/s needs to be maintained in the last through road.

The UCMS simulation program UCMS can be used to simulate cutting sequences. By changing input variables such as shift length, pillar sizes, probability of equipment breakdown, speed of the CM, and speed of the roofbolter, production rates and tramming time values can be obtained. This program was used to simulate the cutting sequences used at Kriel, Greenside, Simunye Shaft, and other developed cutting sequences.

® To end the cutting sequence close to where the next cutting sequence will start.

A combination of these concepts can be used to develop various cutting sequences for Simunye Shaft. The developed cutting sequences can then be simulated to determine which one would be the best for Simunye Shaft.

Developing an optimum cutting sequence According to Shaw (2013), the following factors need to be taken into consideration when a cutting sequence is developed: ® The tramming and cable handling of the CM must be minimized ® Through ventilation must be established as soon as possible ® The roofbolter should not be in the way of the CM ® The tramming routes for all three shuttle cars should be optimized.

Hirschi (2012) conducted a study on Identifying Optimal Mining Sequences for Continuous Miners. In this study he mentions the following guiding policies and practices when developing a cutting sequence: ® Mine crosscuts should be in the direction of ventilation airflow ® A buffer should be maintained between the continuous miner and roofbolter.

Rock mechanical and ventilation standards also need to be taken into account. In terms of rock mechanical standards, not more than three faces may be left unsupported. If three The Journal of The Southern African Institute of Mining and Metallurgy

Results The results of the investigation are presented under the following topics: ® Time study on supporting a 9 m heading at Simunye Shaft. Calculations that follow will be based on this time study ® The effect of geological features on the time to support a 9 m heading. Both the time study and the effect of geological features on support time will be used to set up support targets ® A comparison of the advance rates of the CM and the roofbolter. This will help to determine whether the problem (of RSAT) lies with the roofbolting process ® The effect of implementing hard roof drill bits as a standard product ® Improving the cutting sequence currently employed at Simunye Shaft. Two developed cutting sequences and Greenside and Kriel’s cutting sequences are simulated, and the performance of the four cutting sequences evaluated. A third cutting sequence is developed to improve on the results of the first four cutting sequences.

Time study The time it takes to install a line of support is required to determine whether the roofbolter is too slow and where improvements can be made. If the time taken to install a line of support is known, RSAT logbook data can be used to quantify the benefits of various systems that can be used to reduce RSAT. A summary of a time study, conducted by Van der Merwe (2012), is set out in Table IV. The roofbolt installation sequence, illustrated in Figure 4, was used. The time VOLUME 115

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Figure 10 – Cutting sequence at Kriel Colliery (Odendaal, 2014)

Re-aligning the cutting sequence with general support work

50 6

be set for f situations when such ffeatures are encountered.

As indicated in Table IV, the time to support a 9 m heading can increase from approximately 50 minutes in normal conditions (Table III) to approximately 75 minutes when a feature is encountered.

9

Performance of CM vs. roofbolter

Table IV

Increase in support time when a geological feature is encountered A B C D E F G

Time to support a 9 m heading in normal conditions (min) Lines of support in normal conditions (9 m heading/1.5 m spacing) Lines of support if spacing is reduced to 1 m (9 m heading/1 m spacing) Time to support 9 lines of support (ratio calculation) (min) % increase in time Relocation time (Table III) Roof support advance rate (m/min) (9 m heading/(D+F))

75 50 20 0.095

study was carried out in Ubhejane section at Simunye Shaft.

The bolting took place in normal roof conditions.

A few terms need to be understood to interpret the time study correctly. Total time per row can be described as the time from when the TRS (temporary roof support) is up until just before the TRS is lowered. Table III summarizes the average time it takes to install a roofbolt on each side (LHS or RHS) as well as the total time to support a 9 m heading and the relocation time of the roofbolter. Relocation time is the time from when the roofbolter starts moving to the next heading to the time that the TRS of the roofbolter is up at the new heading, and repositioning time is the time to reposition the bolter to the next hole that needs to be drilled and bolted.

Repositioning time also includes the time to lift and lower the TRS. The average time to insert an extension drill rod is usually around 14–29 seconds (Table III). It was found that the time to insert an extension rod can be extended by almost 2 minutes if a drill bit has to be changed. Drill bit changes can therefore have a major impact on the time to install a line of support.

The changeover to bolting equipment usually takes 49 seconds (Table III), but during the time study it was found that if resin stock runs out it can add approximately 1 minute to the time. Material shortage or poor planning can therefore also contribute to a slower installation time.

The time to support a 9 m heading was determined to be approximately 50 minutes (Table III). Therefore, if a support target has to be set for normal conditions, a time of 50 minutes can be allocated to supporting a 9 m heading.

The effect of geological features on support time If hard roof conditions or slips are encountered, support spacing is reduced, as described in the literature survey. This will result in an increase in support time due to the fact that more roofbolts will have to be installed. The increase in time needs to be established so that a realistic support target can

It is necessary to determine whether the CM or roofbolter advances faster, as this will indicate where the problem of RSAT lies and where improvements can be made. Table V indicates that the CM is faster than the roofbolter in all possible scenarios. The roofbolter can therefore not keep up with the CM in normal conditions. This is a contributing factor to the high RSAT. If the roofbolter advance rate is increased, RSAT may be reduced.

It should be noted that the relocation time (20 minutes) of the roofbolters was taken into account to determine the advance rate. In this report, the advance rate of the roofbolters will be improved to increase the CM advance rate (which already has RSAT intrinsic to it) and reduce RSAT. The increase in the average roof support advance rate will contribute directly to the increase in production.

Solutions to RSAT Eight solutions to the challenge of RSAT were identified, but the most significant solution was to implement hard roof drill bits as a standard product at Simunye Shaft. Only this solution and the use of support targets are discussed in this report.

If normal KCV4 1 RRWT drill bits are replaced with hard roof SV119AE K3012EX02 drill bits (Table II), the life of the drill bits will increase as well as the roofbolt installation rate.

Since the roofbolters at Simunye Shaft are slower than the CMs, the increase in installation rate can contribute directly to reducing RSAT. The hard roof drill bits will prevent operators from continuing to support using KCV4 1 RRWT drill bits in hard roof conditions while they wait for the hard roof drill bits to arrive. Using normal drill bits in hard roof conditions may cause the drill steel and drill bit to expand and become stuck in the adapter, which can result in time wastage.

To determine the benefit of standardizing on hard roof drill bits, the average roof support advance rate has to be determined. For this it is necessary to take all the factors that can reduce the speed of support into account. These include: ® The installation of oslo straps ® The reduction in support spacing when slips are encountered ® Support in hard roof conditions ® Re-support ® Material shortage

Table V

CM and roofbolter advance rates Section

Imvubu Ubhejane Khomonani

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Mining height (m)

Average production rate (t/h)

CM advance rate (m/min)

Roof support advance rate in normal conditions (m/min) *

3.1 3 3

465 322 436

0.23 0.17 0.22

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Roof support advance Roof support rate when slips are advance rate in hard encountered (m/min) roof conditions [Table IV] (m/min) * 0.095 0.095 0.095

0.076 0.076 0.076

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Re-aligning the cutting sequence with general support work Table VI

Table VIII

Roof support advance rate – weighted average (2013)

Increase in average CM advance rate due to an increase in the support advance rate

Normal Multiple slips (oslo strap installation and support spacing reduction) Slips (support spacing reduction) Hard roof Re-support Material shortage Engineering breakdowns Roof support advance rate – weighted average (m/min)

Roof support advance rate in stated condition (m/min)

Weighting (out of 1000)

0.13 0.08

806 40

0.095 0.076 0 0 0 0.11

20 34 20 30 50

C D

Table VII

Improved roof support advance rate – weighted average Condition in which support takes place

Normal Multiple slips (oslo strap installation and support spacing reduction) Slips (support spacing reduction) Hard roof Re-support Material shortage Engineering breakdowns Roof support advance rate – weighted average (m/min)

A B

Roof support advance rate in stated condition (m/min)

Weighting (out of 1000)

0.14 0.09

806 40

0.105 0.082 0 0 0 0.12

20 34 20 30 50

® Engineering breakdowns.

The support advance rate under the conditions specified needs to be determined in order to ultimately calculate the weighted average support advance rate throughout the year (Table VI). The weightings (indicating the ranking of each condition) of the respective support advance rates were estimated from underground logbook data. The weighted average roof support advance rate was calculated to be 0.15 m/min.

When hard roof drill bits are implemented as a standard product, the time to install a roofbolt can be reduced by 30 seconds (Table II). This means that the time to install a line of support can be reduced by one minute. A new support advance rate for each condition can be calculated. The new weighted average support advance rate (Table VII) can be determined by allocating the same weights (as in Table VII) to the respective advance rates.

The increase in the support advance rate can be added directly to the CM advance rate, as the CM advance rate already includes the effect of RSAT. The average CM advance rate for Imvubu, Khomonani, and Ubhejane will increase to 0.221 (Table VIII), which is a 5% improvement in productivity.

As indicated in Table IX, an additional 53 000 t of coal could have been produced in 2013, which is equivalent to 440 hours of RSAT. The result is a 43.13% reduction in The Journal of The Southern African Institute of Mining and Metallurgy

Average CM advance rate in three sections (m/min) Increase in roof support advance rate (m/min) [0.12 - 0.11] % increase in production [B/A * 100] CM advance rate after standardisation (m/min) [A +B]

0.21 0.01 5 0.221

Table IX

Benefit of implementing hard roof drill bits as a standard at Simunye Shaft A B C D E F G H

I J

Average yearly production for 3 sections (Imvubu, Ubhejane, Khomonani) (t) 5% of the average yearly production (tons) [A * 5%] Equivalent RSAT (h) Contribution to reduced RSAT (%) [C/1020] Potential saleable tons Potential revenue (R million) Shifts/year (3 sections) (Mphasha, 2014)1 Average lines of support installed per shift per section (Van der Merwe, 2013) (includes extra bolts for production increase) Additional costs (R per roofbolt) Additional cost per year (R million) [I * 30 * 4 bolts per row *G]2

1 800 000 90 000 440 43.13 53 000 40.12 2050 30

7 1.8

(1) Three production shifts, 5.33 days a week. Therefore: 3 shifts*227days*3 sections ≈2050 shifts (2) A 5% contingency was applied

RSAT. The additional cost per year if hard roof drill bits are implemented as a standard product at Simunye Shaft (for three sections) will amount to R1.8 million. The data in Table IX shows the additional roofbolts that will have to be installed if production is increased annually by 53 000 t for the three sections.

Support targets To increase awareness of the importance of the timeous installation of roofbolts, support targets can be set. Metre targets for the CMs are always visible in the sections and help to formulate goal-orientated tasks for crews underground.

The same effect can be created by setting support targets.

Table X indicates support targets for the No. 4 Seam sections at Simunye Shaft. The different conditions that may arise during roofbolting have to be considered when the targets are set. Therefore, targets for normal conditions, hard roof conditions, and conditions where slips are encountered are indicated in Table X. No cost is associated with this solution.

Cutting sequence development The support challenges arising from the cutting sequence employed at Simunye Shaft can be summarized as follows: ® While the CM is cutting cut number 9 (Figure 11), the roofbolter will be supporting cut number 8 from the left-hand side. This means the roofbolter will be working against ventilation. Workers (roofbolter operators) will be exposed to a lot of dust from the CM and water from the scrubber fan on the CM. This is repeated when the CM cuts numbers 11, 13, 15, 17, and 19. This scenario is illustrated in Figure 11 ® When the roofbolter is supporting cut number 8 from the left-hand side, shuttle cars will be moving in R2 VOLUME 115

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Condition in which support takes place

Re-aligning the cutting sequence with general support work Table X

Support targets, Simunye Shaft Section

Imvubu Ubhejane Khomonani

Metres target per shift

Support target – normal roof conditions (no. of roofbolts)*

Support target – roof with slip (no. of roofbolts )*

Support target– hard roof conditions (no. of roofbolts)*

42 43 42

94 94 94

70 70 70

55 55 55

* Calculated using an effective cutting time of 3 hours per shift and information in Table III

Figure 14 – The re-aligning principle – cutting sequence aligned to support

Figure 12 – Temporary roof support of bolter blocking shuttle car entrance to CM

Figure 13 – Cutting sequence not aligned to support

Figure 15 – Cutting sequence 1

towards the CM. The temporary rooff support off the roofbolter may not be able to reach the end of the split that needs to be supported or may block the shuttle cars that are approaching the CM. This scenario is illustrated in Figure 12.

In response to the challenges presented by the current cutting sequence, the re-aligning principle was developed.

The re-aligning principle refers to aligning the CM cutting sequence with general support work. The concept is explained in Figures 13 and 14. In Figure 13, cuts number 8 and 9 create the challenges. In Figure 14, where the cutting sequence is aligned with support, there is a buffer between the CM and roofbolter. Consecutive cuts (numbers 8 and 9, 10 and 11, 12 and 13, etc.) are further apart. This results in the following advantages: ® Safer condition, because the roofbolter and CM will be further apart (lower risk of collision) and the roofbolter will not have to work against ventilation ® The roofbolter operators will not be exposed to dust and water from the CM ® The CM will not obstruct the path of the roofbolter and RSAT will be reduced.

The following cutting sequences were developed for simulation.

Cutting sequence 1

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As illustrated in Figure 15, boxing takes place in R3 and the cutting sequence ends at the far left-hand side (a lot of tramming time is expected). The re-aligning principle was incorporated into the cutting sequence.

Cutting sequence 2 As illustrated in Figure 16, boxing takes place in R1 and the cutting sequence ends at the far left-hand side (less tramming time than cutting sequence 2 is expected). The realigning principle has been incorporated into the cutting sequence.

It should be noted that both these cutting sequences were approved by the ventilation department as well as the rock engineer at Goedehoop Colliery.

Simulation results In total five cutting sequences were simulated: ® Cutting sequence 1 ® Cutting sequence 2 ® Kriel’s cutting sequence ® Goedehoop’s cutting sequence (the cutting sequence employed at Simunye Shaft) ® Greenside’s cutting sequence.

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Re-aligning the cutting sequence with general support work

Figure 16 – Cutting sequence 2

Figure 18 – Increase in tramming time due to the cutting sequence ending far away from the first cut of the following sequence

Figure 19 – High tramming time as a result of boxing in R3 and ending the sequence far from the first cut of the following sequence

Figure 17 – Simulation results: comparison of the total tramming times of the simulated cutting sequences

The Journal of The Southern African Institute of Mining and Metallurgy

Figure 20 – Low tramming time due to the last cut of the sequence ending close to the first cut of the following sequence

improve on the results obtained, combining all the concepts that resulted in the highest production and lowest tramming time.

Cutting sequence 3: As illustrated in Figure 22, boxing takes place in R3 (through ventilation will be established soon) and the cutting sequence ends close to the start of the following cutting sequence. It can be seen that the re-aligning principle was incorporated into the cutting sequence.

This cutting sequence was approved by the ventilation department as well as the rock engineer at Goedehoop colliery.

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Figure 17 illustrates that cutting sequence 1 had the highest tramming time. It can therefore be confirmed that the greater the distance from the last cut of the sequence to the first cut of the following sequence, the greater the tramming time. Goedehoop’s cutting sequence, which is the cutting sequence employed at Simunye Shaft, had the lowest tramming time. This may be because of the following reasons: ® Goedehoop’s cutting sequence ends close to the first cut of the following sequence (Figure 18 shows the opposite scenario – a sequence ending far from the first cut of the following sequence) ® When boxing takes place in R1 tramming time is added, because the CM has to move all the way from R4 (cut number 11) to R1 (cut number 12), as illustrated in Figure 19, to continue the sequence. This additional tramming time is eliminated if boxing takes place in R3 (Figure 20).

Kriel’s cutting sequence came out top in the comparison of tons booked (Figure 21). A cutting sequence that incorporates the re-aligning principle and has a high enough production output still needs to be developed, as neither cutting sequence 1 nor cutting sequence 2 resulted in a better production output performance than Kriel.

A third cutting sequence was therefore developed to

Re-aligning the cutting sequence with general support work

Figure 23 – Simulation results: comparison of the total tramming time of the simulated cutting sequences (including cutting sequence 3) Figure 21 – Simulation results: comparison of the tons booked for each simulated cutting sequence

Figure 24 – Simulation results: comparison between the total booked tons for the simulated cutting sequences (including cutting sequence 3)

Figure 22 – Cutting sequence 3

Analysis and evaluation of results In this section, the economic viability of the identified solutions are discussed, the simulation results analysed, and the optimum cutting sequence for Simunye Shaft identified by means of a decision matrix.

and average shift production is greater than all the other cutting sequences. Cutting sequence 3 had a 5% greater average shift production than the current cutting sequence employed at Simunye Shaft. This means the mine can potentially earn an additional R40 million in revenue per year.

With any of the cutting sequences that incorporate the realigning principle, the support sequence will be exactly the same as the cutting sequence of the CM, as the roofbolter will be following the CM sequentially.

The results relating to the cutting sequence development and simulation are summarized in Table XI.

Economic viability of solutions to RSAT The costs of mining a ton of coal is set out in Table XII.

Standardizing the drill bits used at Simunye Shaft to hard roof drill bits The break-even analysis (Table XIII) shows that the cost of the hard roof drill bits will be recovered in only 2 months. It should, however, be noted that the cost will be incurred

Table XI

Summary: optimum cutting sequence development Description

Cut 1

Cut 2

Kriel

Greenside

Goedehoop

Cut 3

Average tramming time/ shift (min) Average shift production (t) Is the re-aligning principle incorporated into the sequence?

Is there a buffer between the roofbolter and CM?

29.8

25.9

24.6

25.1

23.7

27.8

2 410.5

2 322.4

2 423

2 374.1

2 394.8

2 432.7

Yes No

x

x x

x

x

Yes No

x

x

x

x

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Re-aligning the cutting sequence with general support work Table XII

Table XIV

Total cost of producing a ton of coal

Weight allocation to cable handling efforts

Description

Cost (R/ton)

Magnitude of effort

Timing

Weight allocated

Mining cost Plant washing cost Rail cost Total

110 60 150 320

High Low

Boxing in R3 Boxing in R1

0 1

Break-even analysis: standardizing to hard roof drill bits A B C D E F G

Potential extra tons produced per year (ROM) Cost per ton (R) Total cost (R million) [A * B] Potential revenue (R million) Potential profit per year (R million) [D - C] Cost of standardizing to hard roof drill bits (R million) [Table IX] Payback period (months) [F/E]

90 000 320 28.8 40.12 11.32 1.8 2

annually. The capital outlay for the hard roof drill bits is low and this is therefore a viable option to implement.

Optimum cutting sequence In order to select the optimum cutting sequence, a decision matrix was set up. The cutting sequences are evaluated against seven criteria, namely: ® Does the sequence incorporate the re-aligning principle? ® Does the sequence result in a low or high tramming time? ® Does the sequence give rise to a high production rate? ® Does the sequence allow the roofbolter to work against ventilation? ® Is there a buffer between the CM and roofbolter? ® Is through ventilation established sooner or later? ® Is the effort of cable handling high or low?

The cutting sequence with the highest score will be recommended. The means of rating the cutting sequences is described to illustrate how the decision matrix was put together.

Does the cutting sequence incorporate the re-aligning principle?

Seeing that the re-aligning principle increases safety and will help reduce RSAT, a high weight has to be attached to it in this selection phase. If the cutting sequence incorporates the re-aligning principle, a score of 5 is allocated, and if not, a 0 is allocated.

Does the sequence result in a high or low tramming time?

Excessive tramming time is inefficient. The higher the tramming time, the lower the potential production. Six cutting sequences were simulated and therefore the sequence with the lowest tramming time is awarded a score of 6 and the sequence with the highest tramming time a 1.

Does the sequence give rise to a high production rate?

Production is directly related to profit. A score of 6 is awarded to the cutting sequence with the highest production output and a 1 to the sequence with the lowest.

The Journal of The Southern African Institute of Mining and Metallurgy

Does the sequence allow the roofbolter to work against ventilation?

Working against ventilation is not good practice and worker safety is enhanced if the roofbolter does not work against ventilation. If the cutting sequence allows the roofbolter to work against ventilation a score of 0 is allocated to the cutting sequence, and if not a 3 is allocated.

Is there a buffer between the CM and the roofbolter?

A buffer between the CM and roofbolter will increase safety due to the fact that roofbolter operators will not be exposed to the dust from the CM and water from the scrubber fan of the CM. If a buffer between the CM and roof bolter is maintained, a score of 4 is allocated to the cutting sequence, and if not, a 0 is allocated.

Is through ventilation established sooner or later?

If through ventilation is established at the start of the sequence, ventilation needs will be met in a more effective manner. If through ventilation is established later, additional fans will have to be installed to maintain an air speed of 1 m/s in the LTR. If through ventilation is established sooner rather than later a score of 2 is allocated to the cutting sequence, and if not, a 0 is allocated.

Is the effort of cable handling high or low?

If cable handling requires less effort, fewer problems can arise to reduce production time. Worker morale will also improve. Cable handling efforts can be divided into categories as indicated in Table XIV. When boxing takes place in R3 the cable has to be suspended all the way to R3 from R1 (where the transformers are), and when the sequence moves towards R1 again, cable handling is doubled. However, if boxing takes place in R1 (where the transformers are) a lot of cable handling effort is eliminated, and flexibility is increased.

The cutting sequence with the highest score is sequence 3 (Table XV). This cutting sequence will ensure safer working conditions, increase production, and reduce RSAT.

Conclusion In completing a root cause analysis of the high RSAT at Simunye Shaft, underground logbook data was analysed and underground workers were interviewed. Seven main contributors to RSAT were identified: machine-related challenges relating to the roofbolter installing support too slowly (the greatest contributor to RSAT), geological conditions (mostly hard roof conditions and slips), logistical challenges pertaining to the CM cutting sequence, manrelated challenges (operator fatigue), re-support, operator inexperience, and the absence of support targets.

A 28% reduction in RSAT was set as the target. The use of hard roof drill bits as a standard at Simunye Shaft was VOLUME 115

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Table XIII

Re-aligning the cutting sequence with general support work Table XV

Decision matrix: optimum cutting sequence Cutting sequence

Does the sequence incorporate the re-aligning principle?

Does the sequence result in a low or high tramming time?

Does the sequence give rise to a high production rate? ventilation?

Does the sequence allow the roofbolter to work against

Is there a buffer between the CM and the roofbolter?

Is through ventilation established sooner or later?

Is the effort of cable handling high or low?

Total score

Cutting sequence 1 Cutting sequence 2 Goedehoop Greenside Kriel Cutting sequence 3

5

2

4

3

4

2

0

24

5

3

1

3

4

0

1

20

0 0 0 5

6 4 5 1

3 2 5 6

0 0 0 3

0 0 0 4

2 0 0 2

0 1 1 0

13 8 13 26

identified f as the best method to address the root cause off high RSAT (slow roofbolt installation). By implementing this solution, RSAT can be reduced by 43.13%, which is more than the 28% target. The total cost of implementing the solution will be R1.8 million per year, with a maximum payback period of 2 months. Simunye Shaft will potentially increase its revenue by R40 million by implementing this solution.

Logistical issues with regards to the CM cutting sequence were also identified as a cause of RSAT, and were thoroughly investigated as suggested by mine management at Simunye Shaft. Challenges with regard to the shaft’s cutting sequence (that may lead to RSAT) were identified and the cutting sequences of Kriel and Greenside Colliery were analysed to improve the situation. A re-aligning principle that increases the buffer between the CM and roofbolter and prevents the roofbolter from working against ventilation, was developed.

Three cutting sequences incorporating the re-aligning principle were developed and simulated together with the current cutting sequence of Simunye Shaft, Kriel, and Greenside Colliery. A trade-off study revealed that cutting sequence 3 had the most promising outcome, and although the effect on RSAT could not be quantified, it was verified that by implementing this cutting sequence Simunye Shaft could increase production by 5%. The support sequence for cutting sequence 3 is equivalent to the cutting sequence itself – the roofbolter can follow the CM sequentially.

Support targets were set up by using the calculated roofbolter advance rate in the various scenarios (hard roof and normal conditions, and when slips are encountered). The advance rate was then multiplied by the effective production time per shift. Findings showed that on average 94 roof bolts have to be installed per shift in normal conditions, 70 bolts when slips are encountered, and 55 when hard roof conditions are encountered. Targets for three scenarios were set, ensuring that the targets are fair and do not demoralise the work force.

Recommendations It is recommended that Simunye Shaft should adopt hard roof drill bits as standard and that support targets for each shift be set. It is also recommended that cutting sequence 3 should be implemented.

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Suggestions for further work The following topics are suggested for further work: ® An investigation into reducing engineering-related RSAT ® An in-depth study of the cable handling logistics surrounding the implementation of cutting sequence 3 in this study ® A study on optimizing logbook data recordings, which will avoid inaccurate data recording ® An in-depth study pertaining to operator fatigue experienced by a roofbolter operators ® Quantifying the effect of changing operators mid-shift, and setting support targets, on RSAT ® A feasibility study on introducing an auto-bolter into sections at Simunye Shaft ® An investigation into change management aspects that need to be taken into account when a new cutting sequence is to be implemented at a colliery.

Acknowledgements I would like to thank L.M. Mphasha, my mentor at the mine, and J.A. Maritz, my supervisor at the University of Pretoria, for their guidance and support.

References BOSCH, C. 2013. Personal communication.

HIRSCHI, J.C. 2012. A Dynamic Programming Approach to Identifying Optimal Mining Sequences for Continuous Miner Coal Production Systems.

Dissertation. Southern Illinois University Carbondale.

MPHASHA, L.M. 2014. Personal communication.

MATHETSA, S. 2013. Personal communication.

MCS. 2013. MCS Roof Bolter Monitoring Proposal for Goedehoop Colliery.

Witbank: MCS. ODENDAAL, A. 2014. Personal communication.

Sinden, J. 2013. Consultant: Fish Eagle Mining Solutions, Personal communication.

STEYN, J. 2010. Introducing the Fletcher Twin Boom, Automated Roof Bolter.

Witbank: J.H. Fletcher & Co. SHAW, E. 2013. Mining Engineer: Khutala Colliery, Personal communication.

STEYN, J. 2013a. Material Handling. J.H. Fletcher & Co., Witbank.

STEYN, J. 2013b. Personal communication.

VAN DER MERWE, B. 2014. Personal communication.

VAN DER MERWE, B. 2013. Personal communication.

VAN DER MERWE, B. 2012. Simunye Shaft Ubhejane section – 9m heading.

Witbank: MCS VAN STADEN, M. 2013. Department of Mining Engineering – Underground Coal Mining Methods. University of Pretoria N The Journal of The Southern African Institute of Mining and Metallurgy

http://dx.doi.org/10.17159/2411-9717/2015/v115n4a2 ISSN:2411-9717/2015/v115/n4/a2

Mining through areas affected by abnormal stress conditions at Syferfontein Colliery

by C. Legote* Paper written as project work carried out in partial fulfilment for BEng (Mining Engineering) degree

This paper investigates the conditions leading to the indefinite termination of production in four critical primary panels at an underground coal mining operation, the observed shortcomings in the mining approach, and the proposed strategy to mine through the affected panels. Initial assessment of the abandoned panel conditions indicated time-dependent strata failure, (i.e. bolted roof failure overrunning intersections), which occurred from mere minutes to up to four weeks post-production, with and without prior warning of failure. This prompted the constant re-supporting of back areas, which raised safety and productivity concerns. Investigation of the initial mining conditions revealed that the failures were due to a critical combination of factors, the chief of which was isolated horizontal stress.

Other factors that were initially overlooked by the mine (i.e. influence of hydraulic stress, misinterpretation of borehole data), resulted in the conditions being described as abnormal. Remedial actions were determined, and in so doing, a new strategic approach was formulated to thoroughly address all failure concerns. The four panels were explicitly planned to serve as the main intake and return airways for the recently commissioned secondary ventilation shaft, as well as providing access to millions of tons in proven coal reserves. It is thus imperative to mine the panels. A feasibility study showed that the proposed strategy set for implementation would be financially viable.

Keywords critical primary panels, time-dependent strata failure, horizontal stress, proven reserves.

Introduction Syferfontein Colliery is an underground coal mining operation situated in Trichardt, Mpumalanga Province, within the Highveld coalfields. At an average depth of 90 m below surface, the mine exploits the 4-Lower coal seam. The bord and pillar method used is fully mechanized, producing on average 2100 t per continuous miner (CM) per shift. Mine expansion required the sinking of a secondary ventilation shaft, the commissioning of which was synchronized with the mining of the four primary panels. The primary panels, once intersected (Figure 1), would serve as the main intake and return airways serving the new ventilation shaft. The panels would also serve as access to millions of tons in proven reserves located in the upper eastern block (Figure 2).

Commissioning of the ventilation was completed, but mining through the panels was halted due to abnormal stress conditions. The nature of the conditions required investigation in order to formulate a strategy to realign the mine with its planned objectives.

The Journal of The Southern African Institute of Mining and Metallurgy

The long-term mine plans were based on numerous factors, chiefly geological in nature.

Syferfontein is riddled with large geological structures including dykes, sills, burnt coal, jointed zones, paleo-lows, and downthrow faults. One such structure is the 13 m wide dolerite dyke that separates the mine’s Riversdale and Weltervreden operations. The focus of this study is on Riversdale.

Study focus Events surrounding the termination of production in the primary panels Both sections were mining concurrently in their respective panels. Section 3 was mining to the north, while section 6 was mining to the east.

Mining parameters were aligned with the standard of 24 m × 24 m pillar centres, 4.1 m mining height, a 7.2 m road width, and advancing 15 m before installing permanent roof support. The factor of safety was determined as 2.28. Omitting the effect of vertical loading, however, a potential high kratio above 2.5 was thought likely (Steenkamp, 2013).

Investigation of the conditions that led to the halt in production within all four primary panels yielded the following results.

Section 3 ® Extensive guttering between pillars, varying in thickness from minor skinning to 45 cm thick chunks, both in supported and unsupported areas within the section, was first observed in panel R31 North Intake, where bolting density was increased from the standard four bolts per row to six, with the row spacing reduced from 2 m to 1 m ® The cutting distance prior to installation of roof support was reduced (sometimes

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Mining through areas affected by abnormal stress conditions

Figure 1 – Detailed locality of panels on mine plan [scale 1:100 000]. Source: Syferfontein survey department (2013)

Figure 2 – Areas of interest (Riversdale): green indicates mined-out and unmineable areas [scale: 1:100 000]. Source: Syferfontein survey department (2013)

to less than 5 m) ® Large volumes of strata water entered the working face, even with the drilled drainage holes ® Observed roof failures were time-dependent, occurring during production and up to four weeks post-production, driving the need to re-support back areas with wire mesh plus additional roofbolts. This had an adverse impact on planned rates of productivity ® Production was first halted in this panel due to the resulting reduction in productivity ® The section moved to the adjacent R30 North Return panel (Figure 3). Similar conditions were anticipated and the increased support strategy was implemented from the onset. Similar failure conditions as in the adjacent panel persisted, but were more serious. These included bolted roof failures that overran intersections, falls of

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ground within splits, and increased guttering between bolts ® As indicated in Figure 3, the nature of the failure conditions drove the section to minimize the number of mining roads from seven to three. This became unproductive and posed further strata and ventilation challenges ® Production in this panel advanced only 250 m from the adjacent panel before it was abandoned due to low productivity and safety concerns.

Section 6 ® Exactly the same failure conditions as in section 3 were observed in section 6. Section 6, however, was mining in an eastern direction more than 1 km from section 3.

Adverse conditions became apparent in the K22 B&P The Journal of The Southern African Institute of Mining and Metallurgy

Mining through areas affected by abnormal stress conditions

Figure 3 – Map indicating conditions in the section 3 panels leading to abandonment of production

® ®

®

®

®

Results Observed shortcomings in the mining approach The normal mining approach failed to address the root causes of the abnormal conditions due to the following shortcomings. ® Long development ends (approx. 15 m) prior to installation of roof support, particularly under laminated/layered roofs, allowing parting that compromised strata competency upon installing permanent support, as emphasized by Steenkamp (2013) ® Trapped water in roof (hydraulic pressure) driving falls of ground even in supported and competent roof areas.

This was aggravated where strata was laminated, as failure under these conditions is violent ® Strata sag or movement was addressed through the installation of telltales. These were, however, only installed in intersections. Installing telltales after roof layers have parted is ineffective. The section telltales were mechanically activated and with the application of stone dust some would be masked and thus overlooked if activated by further sag or movement in the strata.

This made it difficult to control time-dependent failures ® Drainage holes (2 m deep) were drilled only in the The Journal of The Southern African Institute of Mining and Metallurgy

®

®

®

intersections, thus the sources off strata water were not properly identified Dyke structures have a propensity to concentrate stresses around them (Khumalo, 2012). Both sections were approaching a 1 m thick dolerite dyke, but its influence on the stress conditions was overlooked as both section halted production approximately 250 m away from this dyke (Figure 4). The magnitude of concentrated stress is not necessarily related to the size of the feature but rather to the nature and source energy at deposition, as alluded to by Muaka (2013). This was evident with similar cases around the mine, yielding different conditions The impact of isolated horizontal stress was not accounted for in the initial approach. According to Bird ett al. l (2006), the principal horizontal stress direction over the mine region is oriented NW-SE. This conjecture was supported by the observed falls of ground, which were predominantly in a NE-SW direction. Other horizontal stress lead indicators such as buckled bolt plates, reported bumping and spitting in the roof during production, and floor heave (De Clerq, 2013) were noted.

The approach to addressing horizontal stress is not always premeditated as the location and size of the stress source varies throughout the mine The survey cross-section over the two sections’ panels indicated that the immediate strata consisted of competent sandstone. Examination of the rock composition of the failures revealed that the immediate strata in fact consisted of laminated shale and mudstone, which has a major impact on the roof support strategy required Instrumentation used at the time to monitor movements in the strata was by means of mechanical telltales. These were seemingly ineffective due to limited visibility when activated, more so after the application of stone dust The size of the area affected by these abnormal stress conditions cannot be determined due to a lack of detailed VOLUME 115

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East panel, before f the section relocated to the adjacent K23 B&P East panel.

Water-driven falls of ground resulting from bursts of inrushing water occurred before drainage holes could be drilled. Occurrence was close to the face and before roof support could be installed Mining span was also reduced to counteract the falls of ground without increasing the pillar centres.

Similar to section 3, it was found that the failures in all panels followed a NE-SW direction and were aggravated by reactivated joint planes Production was eventually abandoned citing similar reasons as for section 3.

Mining through areas affected by abnormal stress conditions

Figure 4 – Map of section 6 highlighting conditions in the panels

geological data required to prepare the mine stress maps (Van der Westhuizen, 2012).

require the use off battery haulers and a change in cutting sequence, which will pose ventilation constraints.

Suggested strategic approach

Strategy feasibility

The strategy formulated addresses the shortcomings initially overlooked and critical elements that did not form part of the mining COP at the time of failure. Safety took precedence, together with uninterrupted productivity such as to avoid delays caused by time-dependent failures.

A staggered approach with similar dimensions to the normal modified approach (Figure 5) was investigated as it would ideally address the predominant failures in the splits.

The mining direction cannot be altered. This approach would

The feasibility of the proposed strategy was determined using a model designed in conjunction with Sasol Mining personnel to determine the productivity and costs to be incurred until the point of intersection. This model has inputs for mining parameters as well as the required mining consumables, maintenance, and labour.Although the model does not account for operational costs such as electricity, it is nevertheless an effective planning tool. The results are as follows, based on the normal-modified strategy parameters.

Productivity (tons/shift) Mining duration until intersection Extracted coal (tons) Estimated mining OPEX (R mil) *Approx. revenue (R mil) *Avail. working capital (R mil)

Section 3 1100 9 weeks 180 000 2.59 36.00 33.40

Section 6 1000 10 weeks 200 571 2.85 40.11 37.21

*Estimated at a selling price of R200 per ton

Figure 5 – Mining parameters and layout for proposed strategy

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The Journal of The Southern African Institute of Mining and Metallurgy

Mining through areas affected by abnormal stress conditions Table I

Normal-modified layout strategy parameters Normal-modified layout strategy Support requirements

1.8 m long rod, 4 bolts in a row, 1.5 m spacing Breakerline installed around each intersection Rapid installation of support (proactive), wire mesh to act as area support Systematic sidewall support (2 bolts at each pillar corner) W-straps in areas of closely spaced joints , upgrade to electronic 2 m long telltales, proactive drilling of drainage holes 2 m into the roof (not only at intersections as per COP).

Mining parameters

Pillar centres: 24 m × 35 m, bord width 5 m, advance 10 m. Longer pillars in direction of mining, ensure a beam of coal is left beneath the laminated roof (beam to seam height-mining height > 0.6 m minimum). Drill additional inspection holes within roadways during roofbolt installation (not only in intersections) and proactively monitor the strata composition.

Equipment requirements

Refurbished CM with a designated maintenance plan. Two refurbished single-boom roofbolters (addresses high support needs), designated LHD, 3 × 16 t shuttle cars.

Advantages

Adaptable with current cutting sequence, reduced production tempo allows for safer production through adequate time to observe strata behaviour. Ease of ventilation control. Larger pillars account for effects due to horizontal stresses. Strata stability.

Disadvantages

Limited pit room increases equipment congestion thus reducing effective productivity. Rerouting of cables over longer pillars increases relocation time, Routing through ventilation in the LTR will take longer due to reduced advance and longer pillars. Reduced percentage extraction (32.1%) compared to >40% under normal circumstances

Benefits

A reduced road width allows for the creation of stress relief zones, thus less area for stress to act. The creation of two independent teams (possibly from stoneworks teams) will remove the need to use teams and equipment from highproduction sections. Improves stability given the panels will be used as main airways. Using refurbished CM and shuttle cars, which are near their scheduled full overhaul, eliminates the need to buy new machinery for both sections.

Conclusion Production was terminated in four primary panels due to the inability of the mining method to proactively address the abnormal stress conditions, which led to an array of failures.

This had a negative impact on productivity, and was further hampered by ineffective monitoring techniques and misinterpretation of geological data. The abnormal conditions were due to a combination of failure factors that were initially overlooked as a whole. All critical factors have been identified and can be addressed using the proposed strategy. Prior to mining the panels a detailed risk assessment will need to be conducted to ensure that safety standards are aligned. At the completion of the study, the strategy had not yet been implemented, although it was being given strong consideration.

DE CLERQ, A. 2013. Strata control specialist, Rock Engineering Department: Sasol Mining. Personal communication.

KHUMALO, S. 2012. Senior Geologist, Syferfontein Colliery: Sasol Mining.

Personal communication.

MUAKA, J.J.M. 2013. Investigation into the magnitude and direction of ground stresses in the coalfields and their impact on safety and productivity. MSc dissertation, University of the Witwatersrand, Johannesburg, South Africa OOSTUIZEN, P. 2006. Geotechnical aspects of the development of the Sigma Mooikraal underground colliery. ARQ Consulting Engineers (Pty) Ltd. SALAMON, M.D. and MUNRO, A.H. 1967. A study of the strength of coal pillars.

Journal of the South African Institute of Mining and Metallurgy, September 1967. pp. 56–67.

STACEY, T.R. and WESSELOO, J . 1998. In situ stresses in the South African

Recommendations for further work The application of the staggered pillar method for similar conditions and parameters, along with its feasibility in small and large panels as well as the proactive use of extensometers, should be investigated.

mining areas. Journal of the. South Africa. Institute of Mining and Metallurgy, vol. 98, no. 7. pp. 365–368.

STEENKAMP, M. 2013. Strata control specialist, Rock Engineering Department: Sasol Mining. Personal communication VAN DER WESTHUIZEN. 2012. Underground manager: Production Services.

BIRD, P., BEN-AVRAHAM, Z., SCHUBERT, G., ANDEREOLI, M., and VIOLA, G. 2006.

Syferfontein Colliery: Sasol Mining. Personal communication ZHANG, J., LUI, T., ZHANG, Y., PENG, S., and MENG, D. 2002. Strata failure and

Patterns of stress and strain in southern Africa. Journal of Geophysical

mining under surface and groundwater.

Research, vol. III. pp. 1–14.

http://www.link.springer.com/chapter/10.1007/978-3-540-73295-

CAIRNCROSS, B. 2013. Guide to borehole core in the Karoo Basin Coalfields South Africa (1st edn). Struik Nature, Cape Town.

The Journal of The Southern African Institute of Mining and Metallurgy

2_9?no-access=true [Accessed 23 Mar. 2014].

SYFERFONTEIN SURVEY DEPARTMENT. 2013. Sasol Mining.

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References

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http://dx.doi.org/10.17159/2411-9717/2015/v115n4a3 ISSN:2411-9717/2015/v115/n4/a3

A critical evaluation of the water reticulation system at Vlaklaagte Shaft, Goedehoop Colliery

by R. Lombard* Paper written on project work carried out in partial fulfilment of B. Eng. (Mining Engineering)

Water is a very important component in the production process at underground coal mines. Current unfavourable economic conditions have forced the coal mining industry to identify and address every possible bottleneck preventing optimal production. An increase in water-related downtime was identified as one of the bottlenecks at Goedehoop Colliery’s Vlaklaagte Shaft. The purpose of this project was to identify the various causes that contributed to the high downtime (501 hours in 2013, which led to a potential profit loss of R12.9 million) and to suggest possible solutions.

After a thorough investigation the main causes of water-related downtime were identified as low water pressure and low water flow caused by pipe leakages and bursts. The main root cause for the low water flow and pressure was identified as being the low pressure resistance (1600 kPa) of the thin-walled galvanized steel pipes used in the underground inbye water reticulation system. The pipes were selected according to the previous 1000 kPa pressure requirement for the continuous miner.

However, the pressure requirement changed to 1500 kPa, which resulted in the pipes being exposed to much higher pressures than designed for.

The water reticulation system was reviewed and current and future underground pipe layout and water requirements were determined for the shaft. The time frame in which the water consumption would be the highest was determined to be between 1 January 2014 and 7 September 2014. Machine and sprayer specifications were used to determine the water consumption at the shaft.

Three different solutions were considered to solve the water-related downtime problem and to ensure the efficient supply of water to the newly open sections. Permanent underground concrete dams, semi-mobile dams, or new pipe columns with a higher pressure resistance of 3200 kPa were considered. A trade-off study (taking into consideration cost, time to completion and ease of implementation, maintenance requirements, safety, and flexibility) was completed to determine which of these solutions would be most viable.

Keywords water reticulation, down time, pipe bursts, leakages, cascade dam system, permanent dams, portable dams.

Mine background Goedehoop Colliery is situated approximately 40 km east of Witbank in Mpumalanga Province.

Currently Goedehoop has two underground shafts – Vlaklaagte Shaft, which is situated in the southern part, and Simunye Shaft, which is situated in the northern part – which consist of 11 sections. The bord and pillar mining method is employed for coal extraction and each section is equipped with one double-boom Fletcher roofbolter, one feeder breaker, three 20 t shuttle cars, and one continuous miner (CM).

Goedehoop produces 8.7 Mt of run-of-mine (ROM) yearly, of which 5 Mt are saleable.

Ninety-nine per cent of the coal from The Journal of The Southern African Institute of Mining and Metallurgy

Water requirements Water is utilized for many purposes, including dust suppression, cooling, and cleaning (Table I).

Current water reticulation system at Vlaklaagte Shaft As indicated in Figure 1, clean water was supplied to the No. 4 Seam underground sections (via 200 mm galvanized steel pipes) from the surface water cleaning plant until 27 July 2013. The raw water dam received water from the Blesbok reservoir, and the water was then pumped to the water cleaning plant to process the water to drinking quality. However, the pipes that supplied clean water to underground workings from the raw water dam corroded. As a result, recycled water from the RWD (via 200 mm galvanized steel pipes) was used as a substitute. A filtration system consisting of 2 µm sieves was installed to remove solids (which cause blockages in the CM and belt sprayers) from the recycled water.

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Goedehoop is exported through Richards Bay (Becht, 2010).

Vlaklaagte Shaft is currently mining only the No. 4 Seam, as the No. 2 Seam has been mined out. The shaft produced approximately 320 634 t of coal per month in 2013 and made a profit of R120 per ton due to the high-quality coal (on average 27.5 MJ/kg) that is extracted at this shaft (Du Buisson, 2013). The shaft consists of six sections: Section 1 (Simunye), Section 2 (Magwape), Section 3 (Siyaya), Section 4 (Ngwenya), and Block 7 (Section 5/6 and Section 9/10).

The main water source for Vlaklaagte is the Komati Dam. Recycled water from surface is supplied from the return water dam (RWD) to underground sections 1 to 4 via a pipeline running alongside the conveyor belt. Sections 2 and 4 have been developed more than 8 km away from the RWD.

A critical evaluation of the water reticulation system at Vlaklaagte Shaft Table I

Water users and requirements Water users

Requirements

CM sprays

Feeder breaker and conveyor belt sprays

Dust suppression for roads Benicon (mini-pit) Cleaning

Requires water for the following purposes: dust suppression, cooling and cleaning. The CMs operate approximately 9 hours per day. According to Richard Lottering (2013), a Barloworld consultant, CMs requires a flow rate of between 120-135 liters/min and a pressure of 1500 kPa. Failing to adhere to the required flow rate and pressure will result in the CMs tripping which will cause downtime.

Three water sprays are fitted on every feeder breaker for dust suppression. A water spray is also required on every transfer point on the conveyor belt for dust suppression. All the sprays require a water flow rate of 15 liters/min at a recommended pressure of 1500 kPa (Pieterse, 2013) Approximately 60 liters/min is required for road dust suppression (Louw, 2013).

Benicon is a mini-pit near Vlaklaagte that makes use of the water from the RWD and requires approximately 2.1 liters/min.

Cleaning requires approximately 120 liters/min (Horac, 2013)

Since 28 July 2013, water has been supplied to the No. 4 Seam from the RWD. The water cleaning plant therefore only supplies water to the change houses on surface, as recycled water is now being used to supply the underground workings.

at Vlaklaagte. Recently a 150 mm standard galvanized steel pipe size was selected and these pipes were tested to withstand a maximum pressure of 1600 kPa (Louw, 2013).

Summary of water requirements at Vlaklaagte Shaft

Surface pump and pipe layout The surface pump and pipe layout consists of a centrifugal pump (pump 2) which pumps water into a 23 000 litre tank.

The water from the tank is pumped by a five-stage, 65 kW multi-stage pump (pump 1) to the underground sections.

Figure 2 shows the surface pump and pipe layout. Standard 200 mm pipes are used on surface.

Figure 3 is a schematic illustration of the surface to underground pipe layout, including dimensions that are required to calculate the available head.

Underground pipe and pump layout Figure 4 indicates the underground pipe layout and positions of different water users in the different underground sections

Table II is a summary of the water consumption at sections 1, 2, 3, and 4 of Vlaklaagte Shaft (31 December 2013).

Water problems experienced at Vlaklaagte The water-related problems that led to downtime, may be attributed to the following facts. ÂŽ Water is pumped over very large distances, which means that major pipe friction losses need to be overcome. The pressure that is required at the CM has changed over the past years. Previously the CM required only 1000 kPa of pressure to operate. Pipes were selected according to this pressure requirement, and thin-wall galvanized pipes, which can withstand only 1600 kPa, were chosen.

Figure 1—Overview of water reticulation system at Vlaklaagte

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A critical evaluation of the water reticulation system at Vlaklaagte Shaft

Figure 2—Partially flooded suction currently employed at Vlaklaagte

Figure 4—Underground pipe and pump layout at Vlaklaagte Shaft

Figure 3—Surface pipe layout

underground sections, the change from clean water to recycled water, and changes to pump settings and the installation of new pumps) were not well documented

However, the pressure requirement at the CM changed to 1500 kPa, which exposed the pipes to much higher pressures than they were designed for. No action has been taken so far to change the water reticulation system to adapt to this higher pressure requirement ® Vlaklaagte is an old shaft and therefore has an ageing infrastructure, including pipelines. The old infrastructure and increased pump pressures are the main causes of frequent pipe damage and leakages leading to low water flow and low pressure (or no water flow and no pressure) at the face ® Changes made to the water reticulation system over the past years (such as changes in the pipe sizes in

Table III

Downtime hours summary (2013) Section

Downtime hours

Related lost shifts*

50 82 182 187 501

6.3 10.3 22.8 23.4 62.6

Section 1 Section 2 Section 3 Section 4 Total *Note: 8 hours represents 1 shift

Table II

Water requirements at Vlaklaagte Shaft (31 Dec 2013) for Section 1-4 Different activities requiring water

Number of

Flow of water required (l/min)

DOH (hours/day)

Quantity (l/day)

Quantity (l/s)

Quantity (m3/month)

Optimal Pressure required (kPa)

CM (Joy)

2

120

9

129 600

4.0

3 888

1500 in pipe but 2000 at CM

Bucyrus (CAT) Conveyor sprays Feeder Breaker sprays Cars for dust suppression/ fire hydrants* Cleaning (4 sections) Benicon (Mini pit) Total + 10%*

2 25 12

135 15 15

9 21 9

145 800 472 500 97 200 15 000

4.5 6.3 3.0 1.0

4 374 14 175 2 916 450

2

57 600 60 000 107 5470

2.0 0.0 22.9

1 728 1 800 32 264

4

1000-1800 1000-1800 N/A N/A N/A

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*10% was included to compensate for losses

A critical evaluation of the water reticulation system at Vlaklaagte Shaft ® New underground mining blocks, such as the extension in Block 10, for which the current water reticulation was not designed, are being accessed further away from the shaft and the RWD. Data from the water-related downtime logbook was sorted and analysed to determine the extent of the problem and to identify possible root causes leading to the high downtime.

Block 7 (Section 5/6 and 9/10) was excluded from this investigation as Block 7 has a separate water reticulation system in place.

Table III indicates the total hours of production lost by each section from 1 January to 31 December 2013 due to waterrelated downtime. Sections 3 and 4 contributed the most to the total downtime of 501 hours. Solving the problems causing the high downtime in these two sections can eliminate 74% of the water-related downtime. Sections 3 and 4 were therefore selected for further investigations.

A summary of the combined impact of the different causes on both Section 3 and Section 4 is shown in the pie diagram (Figure 5). The chart clearly indicates that low water flow and low water pressure are the two main causes for downtime in these two sections.

total potential profit f lost in 2013 due to water-related downtime was calculated as indicated in Table IV and totalled R12.9 million (Du Buisson, 2013). An intervention was required to stop losses due to water-related problems and to ensure that the water requirements over the life of the shaft are met so that water problems do not occur in the future.

Objectives and methodology The objectives and methodology are presented in Table V.

Results The current water reticulation was reviewed to quantify the reasons for the pipe bursts. The future water reticulation system was also reviewed in order to determine the final pipe layout and underground dam placement.

Analysis of current water reticulation system The pipe layout in Figure 5 can be analysed thoroughly by using the Bernoulli steady-state energy equation (White, 2011):

[1]

Production losses due to downtime Every time production stops the mine loses potential profit. The

Table IV

Water-related downtime cost (1 January 31 December 2013) Section

1 2 3 4 Total

Hours on stop

Cutting rate (tons/hour)

Potential ROM tons

Yield

Sales tons

Potential Profit loss*

50 82 182 187 501

313 323 341 347 1323

15625 26486 61971 64796 168878

0.59 0.61 0.71 0.59 2.50

9219 16156 43999 38229 107604

1.1 1.9 5.3 4.6 12.9

*Potential profit loss = Hours on stop x Yield x Cutting rate x Profit

Table V

Objectives and methodology Objective

Methodology

Quantify the problem

The downtime logbook was thoroughly investigated to: • Determine the total production hours lost due to water-related issues • Determine the potential profit that was lost due to water-related down time • Identify sections with the highest downtime; and • Determine the main causes of the high downtime The company, MCS, was consulted to determine the DOH of the CMs as well as the cutting rate of the CMs. Information on the yield and profit per ton was retrieved by consulting the financial department.

On-site investigations were conducted including: walking the pipelines, observing the different water consumers, manifolds, bends and pumps and where they were located.

The water consumption was calculated by investigating machine and sprayer specifications and also consulting with the Mine Overseer, Shift Boss and Pump Crew at Vlaklaagte Shaft.

The LOM mining plan (obtained from the planning department) for the shaft was investigated and the Mine Planner and Mine Overseer were consulted in order to determine the LOM water requirements.

Information gathered from the mine was used. Various suppliers were also consulted including: • Lectropower • Eljireth • Incledon Recommend the best method for supplying water to the current and newly opened sections.

Review current water reticulation system Investigate and quantify water consumption for the current water reticulation system Determine the life of mine (LOM) water requirements (to prepare for the future) Investigate different methods for supplying water to newly opened sections and solving the water-related downtime problem Draw conclusions and make recommendations from the results of the investigation

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A critical evaluation of the water reticulation system at Vlaklaagte Shaft P2 = Pressure required at the end of the pipe system (at the CM) P1 = Pressure at the inlet V1 = Velocity of the fluid entering the pipe (zero because static water is pumped out of the dam) V2 = Velocity of the fluid required at the end of the pipe system (at the CM) Δz = Height difference/ elevation difference (m).

Equation [2] can be used to correlate the head loss to pipe flow problems (White, 2011).

[2] where f = Friction factor D = Inner diameter (m) K = Minor losses (read off from the table in Appendix H) g = Gravitational acceleration (m/s2) V = Velocity of medium flowing through the pipe (m/s).

Every pipe section has a different flow rate because of the location of the different water users, which results in different frictional losses within each pipe section. A number was allocated to each pipe section in order to differentiate between them (as indicated in Figure 6).

Figure 5—Main causes for water-related downtime in Section 3 and 4

Each term in the equation is a length or a head. α = Kinetic energy correction factor (in problems common to assume that α = 1)

Table VI

Friction factor calculation by using Bernoulli’s equation Section

Length (m)

Component

K factor

Total flow plus 10% wastage (l/s)

u (m/s)

Re

Friction factor*

Hloss (m)**

surface pipe 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 Total

398 100 780 600 1800 700 600 1320 570 910 540 460 50 90 500 420 140 1280 700 140 760 100 100 85 150 325 100 400 285 60

Standard elbow Standard elbow Standard elbow Standard elbow Standard elbow Standard elbow T piece Standard elbow T piece Standard elbow Standard elbow Standard elbow T piece Sharp exit Standard elbow Standard elbow Sharp exit T piece Standard elbow Sharp exit Standard elbow Standard elbow T piece T piece Sharp exit

0 0.45 0.45 0.45 0.45 0.45 0.45 0.9 0.45 0 0.9 0.45 0.45 0

0.0 0.0 9.6 9.4 9.1 8.8 8.5 8.3 0.6 0.3 7.4 0.6 0.3 0.0 6.6 6.3 6.1 2.3 3.3 3.0 2.8 6.3 2.5 2.3 3.3 3.0 2.8 2.5 2.5 2.2

0.0 0.0 0.5 0.5 0.5 0.5 0.5 0.5 0.0 0.0 0.4 0.0 0.0 0.0 0.4 0.4 0.3 0.1 0.2 0.2 0.2 0.4 0.1 0.1 0.2 0.2 0.2 0.1 0.1 0.1

0 0 81 700 79 365 77 031 74 697 72 362 70 028 4 669 2 334 63 025 4 669 2 334 0 56 023 53 688 51 354 19 099 28 011 25 677 23 343 53 688 21 221 19 099 28 011 25 677 23 343 21 008 21 008 18 674

0.01964 0.02157 0.2263 0.0227 0.02277 0.02285 0.02293 0.02302 0.03921 0.04787 0.02331 0.03921 0.04787

0.0 0.0 17.8 1.3 3.7 1.4 1.1 2.3 0.0 0.0 0.8 0.0 0.0 0.0 0.6 0.4 0.1 0.2 0.2 0.0 0.2 0.1 0.0 0.0 0.0 0.1 0.0 0.1 0.1 0.0 30.5

0.45 0.9 1 0.45 0.45 1 0.9 0.45 1 0.45 0.45 0.9 0 0.9 1

0.02366 0.02379 0.02393 0.0282 0.02628 0.02668 0.02714 0.02379 0.02763 0.0282 0.02528 0.02668 0.02714 0.02768 0.02768 0.02832

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*The friction factor was calculated by using the Moody diagram. A friction factor calculator, which can be easily downloaded, was used to accurately determine the friction factor. **The total head loss for each pipe section was calculated by using Equation [2].

A critical evaluation of the water reticulation system at Vlaklaagte Shaft

Figure 7—Future pipe layout

Figure 6—Pipe layout with the numbering of each pipe section (Dec2013)

Table VI details how the friction losses within each pipe section were calculated using Bernoulli’s steady state energy equation. For all the calculations in Table VI it was assumed that e = 0.15 mm and µ = 0.001.

As seen in Table VI the friction losses within the system amount to approximately 31 m. The required head of the pump can now be determined by using Bernoulli’s equation (Equation [1]). Taking into consideration that: ® The static head available (as indicated in Figure 4) is 40 m ® Pressure in the pipes should not exceed 1600 kPa (or 163.2 m) ® The allowable head for the pump can be calculated as 123.2 m (163.2 m – 40 m) ® The frictional head loss in the total length (21 460 m) of pipe is 31 m

® P1 = pgh (h = 2 m, as indicated in Figure 4 the water level in the tank is approximately 2 m above the pipeline exiting the tank) ® P2 = 1500 kPa (the pressure required at the CM is 1500 kPa) ® V1 = 0 m/s ® V2 = 0.13 m/s (derived from the required flow rate of 135 l/min for the Bucyrus CM)

It can therefore be concluded that the pump pressure required for supplying water at the required pressure and flow rate to the four underground sections will cause pipe breaks

Table VII

Future water requirements (section 1, 2, 3, and 4) at Vlaklaagte Shaft (1 Jan 2014 – 7 Sept 2014) Different activities requiring water

Number of

Flow of water required (l/min)

DOH (hours/day)

Quantity (l/day)

Quantity (l/s)

Quantity (m/month)

Optimal Pressure required (kPa)

CM (Joy)

2

120

9

129 600

4.0

3 888

1600 in pipe but 2000 at CM

Bucyrus (CAT) Conveyor sprays Feeder Breaker sprays Cars for dust suppression Cleaning Benicon (Mini pit) Totals Total + 10%

2 33 12

135 15 15

9 21 9

145 800 623 700 97 200

4.5 8.3 3.0

4 374 18 711 2 916

1 600 1 600

15 000

1.0

450

N/A

57 600 60 000 112 8900

2.0 0.0 22.8

1 728 1 800 33 867 37 254

N/A N/A

4

2

*10% was included to compensate for losses

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A critical evaluation of the water reticulation system at Vlaklaagte Shaft Summary of maximum future water consumption at Vlaklaagte Shaft

and bursts. The required pump head (142.21 m) exceeds the allowable head of 123.2 m. No pump will therefore be suitable in this application. Three solutions to this problem were considered: 速 To replace all the thin-walled pipes with thick-walled pipes with a higher pressure-holding capacity 速 An underground cascade dam system using permanent underground dams 速 An underground cascade dam system using semi-mobile underground dams.

The solutions needed to be implemented to satisfy the lifeof-mine (LOM) water requirements. Therefore the LOM pipe layout and maximum future water requirements needed to be determined.

The maximum future water requirement for the shaft was determined to be during the period when sections 2 and 4 moved to block 10 and Section 1 had not been closed yet. A summary of the future water consumption for these four sections is given in Table VII.

Future underground pipe layout The final pipe layout, including final pipe distances for the LOM of Vlaklaagte Shaft, is illustrated in Figure 7. In Figure 8, each pipe section was numbered to facilitate the analysis of the layout.

Table VIII

Pipe friction calculation using Bernoulli's equation Section

Length (m)

surface pipe 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32 33 34 35 36 37 38 39 40 41 42 43 44 45 Total

398 100 780 600 1800 700 600 1320 570 440 200 470 2320 200 900 400 90 1120 440 120 400 540 460 50 90 500 600 420 140 700 140 760 600 100 110 660 200 250 325 100 95 600 90 200 100 60

Component

K factor

Total Flow plus 10% wastage (l/s)

u (m/s)

Re

Friction factor*

Hloss (m)**

standard elbow standard elbow standard elbow standard elbow standard elbow standard elbow t piece standard elbow t piece

0.45 0.45 0.45 0.45 0.45 0.45 0.9 0.45 0.9

19.75 19.75 19.5 19.25 19 18.75 18.5 18.25 7.5 7.25

0.9 0.45 1 0.9

7 3.25 3 3.5

t piece standard elbow Sharp exit standard elbow

0.9 0.45 1 0.45

3.25 3 2.75 0.25

t piece standard elbow standard elbow

0.9 0.45 0.45

5.75 0.5 0.25

t piece

0.9

5

standard elbow standard elbow standard elbow standard elbow t piece t piece

0.45 0.45 0.45 0.45 0.9 0.9

4.75 4.5 4.25 4 3.75 3.5

standard elbow Sharp exit t piece standard elbow standard elbow t piece t piece Sharp exit t piece

0.45 1 0.9 0.45 0.45 0.9 0.9 1 0.9

3.25 3 4.5 4.25 4 3.75 3.5 2.75 0.5

t piece

0.9

0.25

103 729 184 423 182 088 179 754 177 419 175 085 172 750 170 416 70 034 67 699 65 365 30 348 28 014 32 683 30 348 28 014 25 679 2 334 53 693 4 669 2 334 46 689 44 355 42 020 39 686 37 351 35 017 32 683 30 348 28 014 42 020 39 686 37 351 35 017 32 683 25 679 4 669 2 334 -

0.0221 0.02114 0.02116 0.02117 0.02119 0.02121 0.02123 0.02125 0.02302 0.02311

t piece standard elbow sharp exit t piece

0.69 1.23 1.21 1.20 1.18 1.17 1.15 1.14 0.47 0.45 0.44 0.20 0.19 0.22 0.20 0.19 0.17 0.02 0.36 0.03 0.02 0.31 0.30 0.28 0.26 0.25 0.23 0.22 0.20 0.19 0.28 0.26 0.25 0.23 0.22 0.17 0.03 0.02 -

1.43 1.12 8.31 6.24 18.20 6.92 5.78 12.39 0.98 0.71 0.71 0.84 0.06 0.37 0.03 0.35 0.12 0.00 0.57 0.01 0.00 0.40 0.31 0.09 0.42 0.08 0.36 0.25 0.04 0.21 0.14 0.15 0.17 0.05 0.04 0.16 0.00 0.00

0.02321 0.02592 0.02628 0.0256 0.02592 0.02628 0.02668 0.04787 0.02379 0.03921 0.04787 0.02425 0.02443 0.02462 0.02483 0.02506 0.02532 0.0256 0.02592 0.02628 0.02462 0.02483 0.02506 0.02532 0.0256 0.02668 0.03921 0.04787

68.03

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*The friction factor was calculated by using the Moody diagram. A friction factor calculator, which can be easily downloaded, was used to accurately determine the friction factor. **The total head loss for each pipe section was calculated by using Equation [2].

A critical evaluation of the water reticulation system at Vlaklaagte Shaft

Figure 9—Dam placement for leg 1, 2, 3, 4 and 5

Figure 8—Future pipe layout with the numbering of each pipe section

determined by calculating the distances over which the pipe’s maximum pressure rating will be exceeded.

The pipe layout (Figure 7) is too complex to analyse as a single network. The network was therefore divided into five different legs in order to determine how many dams will be required and where the dams need to be placed. The logic behind determining when a dam will be required is simple: the pump needs to supply 153.22 m head at each outlet (spray), but the pipes can only withstand a maximum of 163.43 m, therefore whenever the pump needs to overcome frictional

Analysis of future pipe layout Table VIII shows details of how the friction losses within each pipe section were calculated with the use of Bernoulli’s steadystate energy equation. The total frictional losses were calculated to be approximately 68 m. Table VIII can be used to determine where the underground dams should be placed and how many dams would be required. The placement was

Table IX

Calculation of how many dams will be required in leg 1 and where they are to be placed Section

Friction loss (m)

Pressure required to overcome friction Comment losses and still give the required 153.22 m head at the outlet

Surface pipes 1 2 2-damA damA-3 3 4 4-damB damB-5 damBdamC damC-5 5 5-damD damD-6 6 7 7-damE damE-8 8 8-damF damF-9 9 11 14 16 17 18

1.43

154.65

1.12 8.31 7.66 0.65 6.24 18.2 3.32 14.88 10.21

155.77 164.08 163.43 153.87 160.11 178.31 163.43 168.1 163.43

4.67 6.92 5.54 1.38 5.78 12.39 3.05 9.34 0.98 0.78 0.11 0.71 0.71 0.37 0.03 0.35 0.12

157.89 164.81 163.43 154.6 160.38 172.77 163.43 162.56 163.54 163.43 154.2 154.91 155.62 155.99 156.02 156.37 156.49

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Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

VOLUME 115

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A critical evaluation of the water reticulation system at Vlaklaagte Shaft Table X

Calculation of how many dams will be required in leg 2 and where they are to be placed Section

Friction loss (m)

Pressure required to overcome friction Comment losses and still give the required 153.22 m head at the outlet

Surface pipes 1 2 2-damA damA-3 3 4 4-damB damB-5 damBdamC damC-5 5 5-damD damD-6 6 7 7-damE damE-8 8 8-damF damF-9 9 11 12 13

1.43

154.65

1.12 8.31 7.66 0.65 6.24 18.2 3.32 14.88 10.21

155.77 164.08 163.43 153.87 160.11 178.31 163.43 168.1 163.43

4.67 6.92 5.54 1.38 5.78 12.39 3.05 9.34 0.98 0.78 0.11 0.71 0.71 0.84 0.06

157.89 164.81 163.43 154.6 160.38 172.77 163.43 162.56 163.54 163.43 154.2 154.91 155.62 156.46 156.52

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Table XI

Calculation of how many dams will be required in leg 3 and where they are to be placed Section

Friction loss (m)

Pressure required to overcome friction Comment losses and still give the required 153.22 m head at the outlet

Surface pipes 1 2 2-damA damA-3 3 4 4-damB damB-5 damBdamC damC-5 5 5-damD damD-6 6 7 7-damE damE-8 21 25 25-damG damG-27 27 28 29 30 31 32 34 35

1.43

154.65

1.12 8.31 7.66 0.65 6.24 18.2 3.32 14.88 10.21

155.77 164.08 163.43 153.87 160.11 178.31 163.43 168.1 163.43

4.67 6.92 5.54 1.38 5.78 12.39 3.05 9.34 0.57 0.4 0.3 0.1 0.31 0.09 0.42 0.08 0.36 0.25 0.04 0.21

157.89 164.81 163.43 154.6 160.38 172.77 163.43 162.56 163.13 163.53 163.43 153.32 153.63 153.72 154.14 154.22 154.58 154.83 154.87 155.08

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

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Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

A critical evaluation of the water reticulation system at Vlaklaagte Shaft Table XII

Calculation of how many dams will be required in leg 4 and where they are to be placed Section

Friction loss (m)

Pressure required to overcome friction Comment losses and still give the required 153.22 m head at the outlet

Surface pipes 1 2 2-damA damA-3 3 4 4-damB damB-5 damBdamC damC-5 5 5-damD damD-6 6 7 7-damE damE-8 21 22 23 24

1.43

154.65

1.12 8.31 7.66 0.65 6.24 18.2 3.32 14.88 10.21

155.77 164.08 163.43 153.87 160.11 178.31 163.43 168.1 163.43

4.67 6.92 5.54 1.38 5.78 12.39 3.05 9.34 0.57 0.01 0 0

157.89 164.81 163.43 154.6 160.38 172.77 163.43 162.56 163.13 163.14 163.14 163.14

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Exceeds the maximum 163.43 m that the pipes can withstand - a dam is required

Table XIII

Calculation of how many dams will be required in leg 5 and where they are to be placed Section

Friction loss (m)

Pressure required to overcome friction Comment losses and still give the required 153.22 m head at the outlet

Surface pipes 1 36 37 38 39 40 41 42 44 45

1.43

154.65

1.12 0.14 0.15 0.17 0.05 0.04 0.16 0 0 0

155.77 155.91 156.06 156.23 156.28 156.32 156.48 156.48 156.48 156.48

Table XIV

Weighing of criteria for trade-off study Criterion

Weighting (%)

Cost and payback period Time to completion and ease of implementation Maintenance Safety Flexibility Total

40 20 10 25 5 100

losses exceeding the difference (163.43 m – 153.22 m = 10.21 m), the maximum head that the pipes can handle is reached

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and a dam is required. The calculation for legs 1–5 are presented in Tables IX – XIII. As seen in the tables, seven dams will be required in order to ensure that the maximum pressure of 1600 kPa is not exceeded. The locations of the dams on the underground pipe layout, for all five legs, are shown in Figure 9.

Trade-off study The three possible solutions were traded off, using five criteria: cost, time to completion and ease of implementation, maintenance, safety, and flexibility.

Based on their importance and the preferences of Vlaklaagte Shaft, the criteria were weighted as set out inTable XIV. The solution that scores the highest in the criteria will be recommended for Vlaklaagte.

The Journal of The Southern African Institute of Mining and Metallurgy

A critical evaluation of the water reticulation system at Vlaklaagte Shaft Table XV

Solution 2a

Solution 2b

R 3 875 100

R 438 397

R 2 250 618

It will take a maximum of 1 week to build one U/G permanent dam. This includes transport of the material. Therefore it will take approximately 7 weeks to build the 7 permanent U/G dams. This amounts to 49 days. Eljireth Mining Services are building the U/G dams; therefore the implementation will be very easy for Vlaklaagte, because minimum labour will be required from Vlaklaagte’s side.

It takes Lectropower approximately 3 weeks to build one underground portable dam and to deliver it to the mine. Therefore it will take approximately 21 weeks to build and deliver 7 dams. This amounts to 147 days. Lectropower are building the dams, therefore the implementation will be very easy for Vlaklaagte, because minimum labour will be required from Vlaklaagte’s side.

Low maintenance requirements

Higher maintenance requirements than solution 1. Maintenance of permanent U/G dams is moderate. Vlaklaagte makes use of recycled water U/G and therefore silt will accumulate in the dams. If the silt accumulation becomes too high the dams will have to be cleaned.

Lower maintenance requirements than solution 2a. Maintenance of the semi-mobile U/G dams is less intensive than permanent U/G dams because it has a valve attached to drain the silt if it accumulates.

High safety

If well maintained, high safety. If the dams are well built and maintained there should be no safety hazard.

If well maintained, high safety

Flexible. Most of the pipes can be re-used for other projects after the Vlaklaagte closes.

Poor flexibility. The U/G semi-mobile dams will be not re-usable after Vlaklaagte closes.

Flexible. The semi-mobile U/G dams can be re-used for other projects after Vlaklaagte closes.

Maintenance

Time to completion and ease of implementation

Pipes are installed by Vlaklaagte’s operational team. It takes approximately 1 week to install 1km of pipes, therefore to reinstall 14.5 km length of pipe will take approximately 14.5 weeks, which adds up to 102 days. This includes delivery and transport of the pipes and accessories. The implementation of this solution will be time consuming and more labour intensive that the other two solutions.

Safety

Solution 1

Flexibility

Cost

Summary of how solutions performed against the criteria

Table XVI

Evaluation Solution 1 Weighting factor

100

50

25

0

Total

R1mil-R3mil

R3mil-R7mil

R7mil-R12.9mil

>R12.9mil

20

Cost

40%

Time to completion and ease of implementation

20%

0-1 month to completion. Very easy to implement

2-3 months to completion. Easy to implement.

3-4 months to completion.

Fairly easy to implement.

4-5 months to completion.

Difficult to implement.

>5 months to completion. Very difficult to implement.

10

Maintenance

10%

No maintenance required

Low maintenance

A fair amount of maintenance required

High maintenanceintensive

High maintenanceintensive

7.5

Safety

25%

Completely safe

Very safe

Fairly safe

Low safety

Unsafe

18.8

5%

Completely flexible.

Equipment can be moved around underground with ease and all equipment can be fully re-used after closure of Vlaklaagte

Flexible.

Equipment can be moved around underground with relative ease and some of the equipment can be re-used after closure of Vlaklaagte

Relatively flexible.

Equipment can be moved around underground but with difficulty and very little of the equipment can be re-used after closure of Vlaklaagte

Low flexibility.

Equipment might be moveable underground but with extreme difficulty and very little or none of the equipment can be re-used after closure of Vlaklaagte

Inflexible.

Equipment cannot be moved around underground and none of the equipment can be re-used after closure of Vlaklaagte

3.8

Flexibility

Total

<R1mil

75

100%

Summary off how S h solutions l i performed f d against i the h criteria A summary of how the three solutions performed against the criteria is given in Table XV. This table forms the basis for rating the solutions.

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After f taking Table XV into consideration, the solutions were rated according to the evaluation rubric that was drawn up as indicated in Tables XVI–XVIII. According to the evaluation rubric, building permanent underground dams scored the highest with a value of 73.8.

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Criterion

A critical evaluation of the water reticulation system at Vlaklaagte Shaft Table XVII

Evaluation Solution 2a Criterion

Weighting factor

100

50

25

0

40%

R1mil-R3mil

R3mil-R7mil

R7mil-R12.9mil

Time to completion and ease of implementation

20%

0-1 month to completion. Very easy to implement

2-3 months to completion. Easy to implement.

3-4 months to completion.

Fairly easy to implement.

4-5 months to completion.

Difficult to implement.

>5 months to completion. Very difficult to implement.

15

Maintenance

10%

No maintenance required

Low maintenance

A fair amount of maintenance required

High maintenanceintensive

High maintenanceintensive

5

Safety

25%

Completely safe

Very safe

Fairly safe

Low safety

Unsafe

12.5

5%

Completely flexible.

Equipment can be moved around underground with ease and all equipment can be fully re-used after closure of Vlaklaagte

Flexible.

Equipment can be moved around underground with relative ease and some of the equipment can be re-used after closure of Vlaklaagte

Relatively flexible.

Equipment can be moved around underground but with difficulty and very little of the equipment can be re-used after closure of Vlaklaagte

Low flexibility.

Equipment might be moveable underground but with extreme difficulty and very little or none of the equipment can be re-used after closure of Vlaklaagte

Inflexible.

Equipment cannot be moved around underground and none of the equipment can be re-used after closure of Vlaklaagte

1.3

Total

>R12.9mil

Total

Cost

Flexibility

<R1mil

75

100%

40

73.8

Table XVIII

Evaluation Solution 2b Criterion

Weighting factor

100

50

25

0

Total

>R12.9mil

30

Cost

40%

R1mil-R3mil

R3mil-R7mil

R7mil-R12.9mil

Time to completion and ease of implementation

20%

0-1 month to completion. Very easy to implement

2-3 months to completion. Easy to implement.

3-4 months to completion.

Fairly easy to implement.

4-5 months to completion.

Difficult to implement.

>5 months to completion. Very difficult to implement.

5

Maintenance

10%

No maintenance required

Low maintenance

A fair amount of maintenance required

High maintenanceintensive

High maintenanceintensive

7.5

Safety

25%

Completely safe

Very safe

Fairly safe

Low safety

Unsafe

12.5

5%

Completely flexible.

Equipment can be moved around underground with ease and all equipment can be fully re-used after closure of Vlaklaagte

Flexible.

Equipment can be moved around underground with relative ease and some of the equipment can be re-used after closure of Vlaklaagte

Relatively flexible.

Equipment can be moved around underground but with difficulty and very little of the equipment can be re-used after closure of Vlaklaagte

Low flexibility.

Equipment might be moveable underground but with extreme difficulty and very little or none of the equipment can be re-used after closure of Vlaklaagte

Inflexible.

Equipment cannot be moved around underground and none of the equipment can be re-used after closure of Vlaklaagte

1.5

Flexibility

Total

<R1mil

75

100%

56.5

Conclusions The water-related downtime problem at Vlaklaagte Shaft was quantified through a thorough investigation of the downtime logbook. The main causes of water-related downtime were identified as low water pressure, and low water flow caused by pipe leakages and bursts, the main root cause being the low pressure resistance of the thin-walled galvanized steel pipes used in the underground inbye water reticulation system,

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which cannot withstand the increased pressure now required by the CM. The ageing infrastructure and increased pump pressures are also contributory factors.

The current water reticulation system was reviewed and an underground pipe layout was drawn up for the shaft after onsite investigations. The water consumption of the current water reticulation system was determined from machine and sprayer specifications. The LOM plan was used to determine the The Journal of The Southern African Institute of Mining and Metallurgy

A critical evaluation of the water reticulation system at Vlaklaagte Shaft maximum LOM water requirements, and the time frame f in which the water consumption would be the highest was determined.

Three different solutions were considered to solve the water-related downtime problem and to ensure the efficient supply of water to the newly opened sections. Permanent underground concrete dams, semi-mobile dams, and new pipe columns with a higher pressure resistance of 3200 kPa were considered. The dam placement was determined by calculating the friction loss within each pipe section using Bernoulli’s energy equation. The conclusion was that seven underground dams should be placed to ensure that the maximum pressure of the pipes (1600 kPa) is not exceeded.

The solutions were compared using an evaluation rubric.

Building permanent underground dams was determined to be the cheapest solution (R438 397) and can be implemented in the shortest time (49 days). Cost and time to completion were critical for the solution to be a viable option. The payback period for the cost associated with building underground permanent dams was determined to be 0.035 years, and the solution will save the mine R12.9 million. Building permanent underground dams was therefore identified as the best solution for implementation.

Suggestions for further work

Recommendations

LOTTERING, R. 2013. Consultant, Barloworld. Personal communication.

It is recommended that seven permanent underground dams should be built at Vlaklaagte Shaft to solve the water-related downtime problem and ensure the efficient supply of water to the newly opened sections.

LOUW, N. 2013. Mine Overseer, Goedehoop Colliery. Personal communication.

Acknowledgement I would like to thank Prof. R.C.W. Webber-Youngman, my supervisor at the University of Pretoria, and Charl Du Buisson, my mentor at Goedehoop Colliery, for their guidance and support.

References BECHT, E. 2010. General Manager, Goedehoop Colliery. Presentation.

DU BUISSON, C. 2013. Shaft Manager, Goedehoop Colliery. Personal communication.

HORAC, T. 2013. Foreman, Goedehoop Colliery. Personal communication.

PIETERSE, C. 2013. Section Engineer, Goedehoop Colliery. Personal communication.

WHITE, F.M. 2011. Fluid Mechanics. McGraw-Hill, New York.

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® A sensitivity analysis should be done on the weighting factors of the different criteria used to trade off the three possible solutions. This will give an indication of how changes in the weighting of each criterion would affect the outcome of the trade-off study ® Studies can be done on a more effective recording system for water-related downtime and for recording changes made to the water reticulation system.

Flexibility is the main principle

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http://dx.doi.org/10.17159/2411-9717/2015/v115n4a4 ISSN:2411-9717/2015/v115/n4/a4

Optimization of shuttle car utilization at an underground coal mine

by P.R. Segopolo* Paper written on project work carried out in partial fulfilment of BSc. Eng. (Mining Engineering)

The purpose of the project is to convert current shuttle car utilization on an underground coal mine to best practice by focusing on change-out points and tramming routes, which have a major influence on shuttle car away times. Time studies were an integral part of the project as these enabled the determination of shuttle car away times. An indirect proportional relationship between shuttle car away times and productivity is established. Through the time studies, it is deduced that a third shuttle car will make an insignificant contribution to production when there is only one split open. During this time, maintenance on the third car can be optimized. In order to satisfy the mine’s key performance indicator of keeping shuttle car away times less than 75 seconds, a belt extension must be scheduled after the third split is open. It is established that at any given time, a minimum of two shuttle cars should be used. When cutting on the left-hand-side of the belt road with only two shuttle cars available, the centre and left (left of the feeder breaker) shuttle cars should be used for coal hauling. When cutting on the right-hand-side, the centre and the right-hand-side cars should be used. If only one shuttle car is available, the centre car is the most efficient to use. Alternative anchoring configurations can be employed so as to enable cars (left or right, especially) to reach the opposite extremities of the panel and hence minimize cable length restrictions.

Keywords coal mining, underground transport, coal hauling, tramming, scheduling, optimization.

Introduction Zibulo, meaning g first born in Zulu, is the first new mine in the Anglo American Inyosi Coal (AAIC) joint venture. It was formerly known as the Zondagsfontein coal project. The project is majority-owned (73%) by Anglo American and the remaining 27% by Inyosi, the black economic empowerment company formed in 2007 as part of Anglo Coal’s second wave of empowerment in South Africa (Anglo American, 2007). The colliery is situated in Ogies in Mpumalanga Province. With a life of mine of 20 years, the project comprises of two operations; an opencast and an underground operation. This project was carried out at the underground operation.

Zibulo Colliery is sited within the Witbank Coalfields, which are usually comprised of five seams numbered (from the base upwards) No. 1 to No. 5 seam. The colliery extracts No. 2 seam. The disturbed and relatively shallow The Journal of The Southern African Institute of Mining and Metallurgy

Project background Coal hauling background at Zibulo Coal is hauled by means of both battery haulers (BHs) and shuttle cars (SCs) at Zibulo Colliery’s underground operation. The background of the coal hauling equipment at the eight sections of the mine is illustrated in Figure 1.

A total of 21 new coal haulers were initially purchased; 9 SCs and 12BHs. The SCs were employed in sections 1, 2, and 3.

Sections 4, 5, 6, and 7 were using BHs. During this start-up phase, three redundant BHs were purchased from Goedehoop colliery, an Anglo American Thermal Coal underground operation. These machines had to be overhauled in order to get them into operational condition; they were then put into production in section 8 as a temporary solution (Anglo American, 2013). In 2013, they were deemed to have reached the end of their life cycle. They displayed low availabilities and thus had to be replaced with new SCs (Anglo American, 2013). Six new SCs came on stream in 2013. Sections 4 and 6 are currently using three SCs each. The BHs from sections 4 and 6, however, were split among sections 5, 7, and 8. The initial BHs from section 8 are no longer in production; section 8 currently uses the three BHs from section 4 and an additional hauler from section 6. A BH from section 6 was added to the three haulers in section 5.

The other hauler from section 6 (added to section 7’s fleet) is not currently operated.

* University of the Witwatersrand, Johannesburg, South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper rreceived Feb. 2015 VOLUME 115

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Synopsis

(depth approximately 100 m) coal seam is mined using the bord and pillar method due to the low capital investment and operating costs required, together with its level of selectivity and safety.

Optimization of shuttle car utilization at an underground coal mine

Figure 1—Background of coal hauling equipment at Zibulo

It is clear that SCs have found f favour f at the operation.

Zibulo Colliery currently employs a total of 15 Joy 10SC2256C machines in five (sections 1, 2, 3, 4, and 6) of its eight sections. The SCs have the following specifications (Anglo American, 2013).

Supplier Joy Mining Machinery Drive Dual conveyor motor drive Control VFD OPTIDRIVE traction system Pump capacity 25 kW pump motor Lubrication Auto lubrication system Traction power 2 × 85 kW traction motors Capacity 20 t Minimum seam height 1.96 m

20 m off the total SC cable slack is from f the anchor point to the switches, which are placed parallel to the FB. The points where SCs interact with each other (indicated by squares 1 and 2 in Figure 4) are the change-out points. At any one of these points, a SC waits on the next before it proceeds towards the CM to avoid running over the next SC’s cable. At square 1 in Figure 4, the red car waits on the blue car.

Cutting sequence and roads A typical cutting sequence from section 4 is illustrated in Figure 2. The road on which the feeder breaker (FB) is located is referred to as the belt road (BR). Roads on the left of the BR are the left roads (referred to as L1, L2, L3 to the further left of the belt road), to the right of the BR, roads are R1, R2, R3, R4. This particular sequence is characterized by eight roads; other sections have a different number of roads, depending on the panel width. Ventilation in a section is directed from right to left; it is for this reason that cutting in each section generally takes place from the right to the left of the section. Once the continuous miner (CM) has made its first cut between R3 and R4 it is trammed to the second cut on R3, thereby making way for roofbolting to be carried out where the first cut took place.

Figure 2—Typical section 4 cutting sequence and roads

Through roads Through roads are also referred to as splits; these are illustrated in Figure 3. These are open roads between the FB and faces to be cut. Zibulo Colliery maintains a maximum of three through roads in each section. When three splits are open, the belt is extended (i.e. the FB moves towards the faces) two splits ahead.

Shuttle car change-out points, tramming routes, anchor points, and switches The green, blue, and red circles in Figure 4 represent three shuttle cars in a section. Their respective dashed lines represent their trailing cables and thus the way in which the shuttle cars tram towards and back from the CM at R3. These cables are anchored on the three points adjacent to the FB, where the SCs all tip from their three distinct points. About

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Figure 3—Through roads The Journal of The Southern African Institute of Mining and Metallurgy

Optimization of shuttle car utilization at an underground coal mine the section as well as their tramming routes was established.

On a few occasions, the time it took for a SC to tip onto the FB was recorded. The results obtained were predominantly from direct time measurements. From these, the tramming distances and average tramming speed were obtained.

As a means of confirming the consistency and reliability of the results, they were compared to the Joy CM system that the mine uses to monitor whether the set production KPIs are being met. Although a great degree of similarity was observed from the comparison, data calculated from the tramming distances and the average tramming speed could not be confirmed. In an attempt to clear this hurdle, the Arena simulation software program was used. The program was able to confirm the findings from the studies.

Figure 4—Shuttle car change-out points, tramming routes, anchor points, and switches

Shuttle car away times Problem statement and aim of project Each section is equipped with a FB on which coal can be tipped from three distinct points. The mine took advantage of this by employing three SCs (each anchored at a distinct point) in each of five sections, with the aim of maximizing productivity. However, the overall SC utilization has decreased from 2010 to 2013. Furthermore, data gathered from CM operational reports reveals that CM waiting times (or SC away times) in all five sections employing SCs for coal haulage are rather long compared to the mine’s key performance indicators (KPIs). The decrease in SC utilization and the longer SC away times leads to lower production rates.

The project is therefore aimed at increasing productivity through the optimization of SC utilization.

Results and analysis In each of the five SC sections, a considerable amount of time was spent near a CM recording SC away times as well as SC loading times. Away time in this study was taken as the time between completion of loading one SC and the arrival of next (or the same car if only one is being used) to be loaded.

Before each recording session (or each time the CM had to tram to another cut) the numbers of SCs being operated in

From observations, SC tramming routes were similar in all five sections. Long tramming routes lead to longer away times; longer away times lead to lower production rates. The inverse relationship between the tramming route distances and productivity is thus established. The main objective, therefore, is to keep tramming routes as short as practically possible with respect to the CM position, CM cables, and ventilation.

Tables I–IV represent the average away times that were obtained from the data. The mining height is approximately 3.5 m; the pillar and bord widths are 12 m and 7.2 m respectively. Approximate distances between the FB and CM, between the FB and the main change-out point (COP), and from the COP to the CM were determined using these bord and pillar widths. In Tables I, II, and III, S/C columns denote the following: ® 1 S/C: SC X ® 2 S/C: SCs X and Y ® 3 S/C: all three SCs (X, Y, and Z).

For the columns ‘FB to CM’, ‘FB to main COP’, and ‘COP to CM’: ® 1st S/C: SC X ® 2nd S/C: SC Y ® 3rd S/C: SC Z

Table I

Average away times for one split Average away time (seconds)

Distances (metres) FB to CM

FB to main COP

COP to CM

Road

1 S/C

2 S/C

3 S/C

1st S/C

2nd S/C

3rd S/C

2nd S/C

3rd S/C

2nd or 3rd S/C

1 2 3 4 5 6 7 8 9 10 11 12 14 16 18

R3 & R4 R3 R4 R2 & R3 R4 R3 R1 & R2 R3 BR & R1 R2 R1 R2 R1 BR BR

114 115 128 117 137 126 128 129 115 111 88 109 121 73 82

25 26 35 45 40 31 34 35 31 20 17 25 26 21 28

26 25 35 44 42 31 35 35 30 20 17 26 26 17 26

99 99 118 137 130 111 118 117 99 80 61 92 73 80 92

99 99 118 147 130 111 118 117 99 80 61 92 73 42 54

137 137 156 175 168 149 156 155 158 139 120 151 111 80 92

78 78 78 78 78 78 78 78 59 59 57 59 40 21 21

116 116 116 116 116 116 116 116 97 97 116 97 78 59 59

21 21 40 59 52 33 40 39 40 21 21 33 33 21 33

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Cut

Optimization of shuttle car utilization at an underground coal mine Table II

Average away times for two split Average away time (seconds)

Distances (metres) FB to CM

FB to main COP

COP to CM

Cut

Road

1 S/C

2 S/C

3 S/C

1st S/C

2nd S/C

3rd S/C

2nd S/C

3rd S/C

2nd or 3rd S/C

3 5 1 2 6 8 4 10 12 7 11 14 9 18 16

R4 R4 R3 & R4 R3 R3 R3 R2 & R3 R2 R2 R1 & R2 R1 R1 BR & R1 BR BR

146 155 132 133 142 147 163 118 126 146 102 111 131 97 89

46 49 37 37 41 45 52 30 34 45 22 29 46 36 29

36 43 26 26 30 37 50 21 31 36 21 23 44 25 17

139 151 120 120 132 141 158 101 113 139 82 94 120 113 101

139 151 120 120 132 141 158 101 113 139 82 94 120 75 63

177 189 158 158 170 179 196 139 151 177 120 132 158 113 101

99 99 99 99 99 99 99 80 80 80 61 61 61 42 42

137 137 137 137 137 137 137 118 118 118 99 99 99 80 80

40 52 21 21 33 42 59 21 33 59 21 33 59 33 21

Table III

Average away times for three split Average away time (seconds)

Distances (metres) FB to CM

FB to main COP

COP to CM

Cut

Road

1 S/C

2 S/C

3 S/C

1st S/C

2nd S/C

3rd S/C

2nd S/C

3rd S/C

2nd or 3rd S/C

1 2 3 4 5 6 7 8 9 10 11 12 14 16 18

R3 & R4 R3 R4 R2 & R3 R4 R3 R1 & R2 R3 BR & R1 R2 R1 R2 R1 BR BR

147 147 162 177 171 157 162 164 147 133 118 142 127 104 113

45 45 52 59 57 49 52 53 45 38 30 42 35 37 42

21 21 37 51 44 31 37 37 36 21 21 31 31 16 25

141 141 160 179 172 153 160 162 141 122 103 134 115 84 96

141 141 160 179 172 153 160 162 141 122 103 134 115 122 134

179 179 198 217 210 191 198 200 179 160 103 172 153 122 134

120 120 120 120 120 120 120 120 101 101 82 101 82 101 101

158 158 158 158 158 158 158 158 139 139 120 139 120 101 101

21 21 40 59 52 33 40 42 40 21 21 33 33 21 33

Table IV

Calculated away times for four split Calculated away time (seconds)

Distances (metres) FB to CM

FB to main COP

COP to CM

Cut

Road

1 S/C

2 S/C

3 S/C

1st S/C

2nd S/C

3rd S/C

2nd S/C

3rd S/C

2nd or 3rd S/C

1 2 3 4 5 6 7 8 9 10 11 12 14 16 18

R3 & R4 R3 R4 R2 & R3 R4 R3 R1 & R2 R3 BR & R1 R2 R1 R2 R1 BR BR

164 164 178 199 187 173 178 180 164 149 134 158 144 149 158

53 53 57 67 65 56 60 61 53 31 38 50 41 46 50

21 21 37 51 45 31 36 37 36 21 21 31 31 21 21

162 162 181 200 193 174 181 183 162 143 124 155 136 143 155

162 162 181 200 193 174 181 183 162 143 124 155 136 105 117

200 200 219 238 231 212 219 221 200 181 162 193 174 143 155

141 141 141 141 141 141 141 141 122 122 103 122 103 84 84

179 179 179 179 179 179 179 179 160 160 141 160 141 122 122

21 21 40 59 52 33 40 42 40 42 40 21 33 21 33

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Optimization of shuttle car utilization at an underground coal mine The different ff routes to be trammed by the SCs are illustrated in Figure 5. Note that cycle time study results are tabulated for only the belt road (BR) to the R4, and not from the BR to the left-hand-side extremities of the panel. This is because mirrored results for cuts on the roads would be obtained for the BR to the L3 such that away times from the 2nd cut would be the same as that of the 22nd cut; away times from the 14th cut would be the same as that of the 21st cut. Mining is carried out from the right-hand extremities of the panel to the left for ventilation purposes.

Tramming routes for cuts 9, 6, and 8 are illustrated in Figure 5. These are shown by the dashed lines with the respective shuttle car colours. The main COPs are indicated by the transparent spheres, at these points, all three SCs interact with each other. Note that there is no distance, for there is no change-out-point when the no. 1 SC is operated solely; it does not interact with the next car.

Zibulo Mine standards allow for a maximum of only three splits before a belt extension. Although contraventions of the standard were not observed in practice, away times when there were four splits between the FB and the face were not directly recorded. These were calculated from the data already obtained. The average speed at which SCs tram was determined; the tramming route distances to each cut were computed. Away times were then determined from dividing the route distances by the SC average tramming speed. The determination of away times when using two or three SCs, however, was rather complex. This involved manual simulations that were only carried out on paper.

With maximum SC cables lengths of approximately 230 m, it is evident from the table that SCs may not be able to reach certain cuts from their anchor points. The 3rd SC would not be able to reach cuts 1 to 9 as well as cut 12 because the maximum SC cable lengths are specified in the mine standards.

not change when a third car is added to the two that are already being operated.

A similar trend is observed from Figures 7 and 8. Adding a second car considerably increases production rates, as it can be observed that away times decrease. A third SC adds value at splits 2, 3, and 4. In the case of a single split, the average away times between SCs 2 and 3 does not change.

This suggests that adding a third SC when two are already operating in the section will not increase the production rate.

It can therefore be deduced that maximum practical production can be achieved by using only two SCs, when there is a single split between the FB and the face being cut.

Figure 6—Average away times when cutting the second cut at distinct splits using 1, 2, and 3 shuttle cars

Analysis of shuttle car away times

Figure 7—Average away times when cutting the 7th cut at distinct splits using 1, 2, and 3 shuttle cars

Figure 5—Different routes to be trammed by the shuttle cars

Figure 8—Average away times when cutting the 11th cut at distinct splits using 1, 2, and 3 shuttle cars

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As a means of analysis, three random cuts were selected. The effects of the number of SCs being operated for a number of splits were analysed from Figure 6, 7, and 8. A general trend from all three graphs was observed.

When using a second SC as opposed to just one, away times from all four split numbers decrease significantly. A second SC therefore adds significant value; a production rate increase is realized when using two cars rather than one.

When a third SC is added, however, away times from splits 2, 3, and 4 decrease. From Figure 6, away times with split do

Optimization of shuttle car utilization at an underground coal mine It is to be noted, on the other hand, that the average away times in the case of a single split decrease when a third SC is added to the operation only if the CM is cutting at the belt road. Average away time decreases from 21 seconds to 17 seconds when cutting the 16th cut by adding on a third SC at the first split. Therefore, a third SC adds production value only when the CM is cutting at the BR (cuts 16 and 18).

To check if the deduction is valid, Arena (mining simulation program) demonstrations were made available.

According to Olivier (2014), ’the shuttle car on the right takes up two thirds of every 3 shuttle car loads when the CM is cutting on the right-hand-side of the belt road’. When the CM is cutting on the belt road, each of the three cars takes up a third of every three SC loads. Olivier’s statement thus confirms that a third SC is rather insignificant when there is a single split between the FB and the face.

these away times implied a 3.48% production loss from f change-out point A to B. Siyanqoba section (section 4) had a production target of 1 Mt for 2014. Designing for change-out point B would lead to a production loss of only 34 800 t; the section may only produce 965 200 t. A 3.48% production loss is not of great consequence if it promotes safety.

Shuttle car configurations Since each SC section employs three SCs, it has become generally accepted that all three cars should be running at all times in order to meet production requirements. Not only is this not necessarily the case, but utilizing all three cars at all times is somewhat impractical, due to factors including cable lengths and operator availabilities. It therefore becomes necessary to factor in SC configurations that will lead to maximum productivity. These should be used at all times if production targets are to be met.

Safety implications At Zibulo colliery, safety is a core value. It is common practice to have the change-out points as close as possible to the face being cut. Figure 9 illustrates change-out points A and B when the CM is cutting the 12th cut. The stars on both diagrams represent where the CM operators would stand during operation. Having a change-out point directly adjacent to the CM operators’ position brings about the risk of a SC running into the operators as it attempts to make a tight turn to position itself in place for loading. For safety reasons, designing the change-out point to be at A is therefore not recommended. With the common practice borne in mind, it was deemed necessary to carry out a further investigation on the implications this may have on safety in relation to production.

Time studies of the away time differences between change-out points A and B were conducted. The difference in

Optimal car configurations determined from simulations Table V highlights the effects of SC configurations on production, utilizing simulations run on the Underground Coal Mining Simulation (UCMS) program. Different numbers of cars and their configurations were simulated to run throughout the entire panel length from the 1st to the 68th cut, illustrated in Figure 10. The cutting sequence in Figure 10 was initially input as Zibulo’s cutting sequence into the program. It is important to note, however, that this cutting sequence varies from that generally followed at Zibulo Colliery’s underground sections. According to Olivier (2014), the data obtained from the program can still be relied on as cars are trammed to all different cut locations, only in a different sequence.

Figure 9—Different possible change-out points when excavating cut 12

Table V

The effect of shuttle car configurations on productivity 3SC

No RHS

No centre

n=No LHS

Only centre

Only RHS

Only LHS 238.11

Cycle time

114.79

128.73

131.01

128.54

197.72

234.78

Av. production rate (TPH)

556.01

496.82

488.27

497.35

324.23

274.32

271.37

Production time (min)

280.84

272.87

274.65

272.65

310.24

322.63

322.82

2259.47

2235.01

2260.07

1676.46

1475.04

1460.05

-13%

-14%

-13%

-36%

-43%

-44%

2

2

2

1

1

1

Tons (booked)

2602.5 0%

No of Cars

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Optimization of shuttle car utilization at an underground coal mine different ff numbers off cars running in the section. A 76% production improvement can be obtained when using three cars instead of just one. Employing two SCs instead of only one offers a 55% production improvement. The least production improvement is achieved when transitioning from using only two SCs to three; this offers a 15% production improvement. Although using all three SCs offers the greatest practical productivity, using two (even though to a lesser extent) cars is also viable.

Applicable shuttle car configurations determined from away time studies It is one thing to understand the value that different car configurations add to production; it is another to determine when and which SCs to apply when extracting coal at any particular cut. SC average away times when the CM cuts at seven distinct points, illustrated in Figure 12, were analysed so as to identify which car configurations to apply.

It has already been established that the effect of a third SC is rather insignificant when there is a single split between the FB and cut. With two splits, however, using all three cars offers maximum production. Using two SCs is the secondbest option. Table VI shows the SC average away times when

Figure 10—Zibulo Colliery's cutting sequence according to UCMS

According to UCMS, 2 602.5 tons can be booked when all three SCs run for a production time duration of approximately 4 hours and 41 minutes. The benchmark was therefore assumed to be when running all three SCs, hence the zero production decrease indicated in Table V. Variations on this were then investigated in order to see how the results differ from the benchmark.

Using only the centre car together with either the right or left SC results in a 13% production decrease. When using both the right and left cars with no centre car, UCMS suggests that the tons booked will decrease by 14%. It does happen that a section has only one SC running at a given time due to maintenance, breakdowns, or operator availabilities. This case was also investigated. According to UCMS, a maximum of 44% of the benchmark tons booked can be lost when running only one SC. This is when the left-hand side (LHS) car is solely used. The least production loss when employing only one car is obtained when only the centre car is used.

From the 1st cut all the way to the 68th, all six configurations are applied (or rather as a result of breakdowns, car maintenance, or labour and cable management) at any given time.

Figure 11—Summary of production improvement when using different numbers of shuttle cars

Production improvements offered by different car configurations As summarized by UCMS, the maximum improvement in productivity is achieved when employing all three SCs instead of only one. This is illustrated in Figure 11; a summary of the improvement in production when comparing the usage of

Figure 12—Cuts of interest in a typical section 4 cutting sequence

Table VI

Average away times for six shuttle car configurations when being loaded at seven distinct cuts

5 10 9 18 19 15 22

Average away times Road

RHS only

LHS only

Centre only

RHS+centre

LHS+centre

LHS+RHS

R4 R2 BR/R1 BR L1 L2/L1 L3

155 117 131 126 161 161 190

184 146 161 126 102 131 131

155 117 117 97 102 102 131

49 30 46 34 52 52 66

71 67 49 34 22 37 37

112 44 51 34 57 59 68

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Optimization of shuttle car utilization at an underground coal mine cuts 5, 9, 10, 15, 18, 19, and 22 are being excavated. It can be seen that minimum average away times are obtained when using the right-hand side (RHS) car together with the centre one; this is when the cuts on the RHS of the BR) are being mined. Average away times when using the LHS and centre cars are somewhat moderate; those of using both the LHS and RHS cars are the highest. This suggests that when cutting on the RHS of the BR, a configuration (involving the use of two SCs only) of the RHS and centre car is the most viable for meeting production targets.

When excavating cut 18, which is positioned on the BR, any pair of SCs is viable. This is due to the constant average away time (34 seconds) offered by any pair. When cutting on the LHS of the BR (cuts 19, 15, and 22), on the other hand, minimum average away times are achieved when using the LHS and centre SCs. Again, the average away times when employing both the RHS and the LHS cars are the highest.

It is not always possible to have either two or three SCs running at any given time in a section. It is therefore necessary to determine which single car should be employed, as well as when it should be applied. When cutting on the RHS roads of the BR, minimum average away times are achieved when using either the RHS or the centre cars, as both their average away times are equal. When cutting on the RHS split of the BR, such as cut 9, minimum average away times are obtained when using only the centre car. This is attributed to the fact that SC tramming routes when cutting at this point are not as straightforward as when cutting on the roads. As expected, the least average away times when cutting at the BR are achieved when using the centre car only. Using either the LHS or centre car offers minimum average away times when cutting on the LHS roads of the BR. Similarly, when cutting on the LHS splits of the BR, such as cut 15, the least average away times are achieved when the centre car only is operating. When cutting at this point, using the RHS car only offers the highest average away times.

Challenges and opportunities Of course, the utilization of SCs in most of Zibulo colliery’s underground sections presents more advantages than the cars’ counterparts, the battery haulers. However, as De Lange (1988, p. 151) previously postulated about the future of underground transport on large coal mines, ‘coal mining is a major transport business and hence there will always be new challenges to meet in underground transport’. SC operations

involve the consideration off various ffactors, both technical and non-technical. A challenge presents an opportunity to improve or employ new techniques. The challenges of SC utilization and the associated opportunities they present are discussed in the following sections.

Tramming routes A direct relationship between the tramming route distances and the average away times has already been established; it is therefore important to keep the tramming routes as short as practically possible. Not only should the distances be kept minimal; the following factors should be considered when designing or determining tramming routes.

Avoiding turns as far as practically possible Figure 13 illustrates the possible tramming routes that the centre SC (SC X) can follow when the CM is cutting at the 10th cut. The tramming route illustrated on the LHS diagram of Figure 13 shows that car X would have to make four turns to and from the CM to the FB . As stated by Smit (2014), ‘a tramming route should have as little turns as possible’.

Tramming routes with more turns are both unsafe and ineffective, such that car X on the LHS diagram will have longer average away times than car X on the RHS diagram of Figure 13 (Smit, 2014). Like any other trackless equipment, the tramming (or hauling) speed is reduced when turning. On the LHS diagram, the average shuttle car speed of 2.7 m/s (obtained from SC time studies) will be reduced. As the average away times consequently increase, the production rate decreases.

For any equipment, although training is offered, ease of operation should be the main objective. A tramming route with more turns compromises the ease of SC operation. The cars can easily collide with pillars. An operator may become fatigued relatively quicker compared with the operator operating car X on the RHS diagram.

Tramming route obstructions If the shortest tramming routes are to be used at all times, it is important to make sure that they are free of any obstructions. Figure 14 shows a SC cable and a brattice that were placed on the area indicated by the star on the RHS diagram. This implies that car X used the tramming route illustrated on the LHS diagram of Figure 13. This constitutes good housekeeping as well as thorough and effective planning. Brattices should be installed such that they do not obstruct the desired tramming routes. For example, the

Figure 13—Different tramming routes for shuttle car X when the CM is excavating cut 10

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Optimization of shuttle car utilization at an underground coal mine

Figure 14—A shuttle car cable and brattice placed in the section

installation off the brattice indicted by the dashed line on the RHS diagram of Figure 14 should be delayed until cuts 3, 4, and 5 are excavated so as to clear the shortest, most practically possible and viable tramming routes to these points.

Floor conditions Bad floor conditions can be as a result of an uneven floor (attributed to geological conditions or floors that are not swept), poor water drainage, and steep gradients. These conditions may significantly reduce the life of the cars’ components and consequently cause premature failure, or the cost per ton of the operation may increase due to losses in efficiency and productivity (Callow, 2006, p. 821). At Zibulo Colliery’s underground sections, coal extraction is carried out on a relatively flat gradient. Tramming routes are relatively flat and so shuttle cars tramming at high gradients is not a concern. The section floors are generally kept in good condition. During the rare cases of floor flooding, however, corrective measures to drain water should be taken as quickly and efficiently as practically possible.

two SCs. This then becomes a cycle; the belt is extended two splits ahead with one split open between the FB and line of cut, the 2nd split is open and as soon as the 3rd is entirely open another belt extension should be carried out. The initial 3rd split will therefore become the 1st split after the belt is extended.

Cable management At Zibulo Colliery, cable management is a major concern. The mine is relatively new and so the effective management

Belt extension To maintain overall short tramming distances, it is important to schedule a belt extension effectively. This means that there should be a maximum number of splits open before each belt extension. To determine this, the overall average away times when using all three SCs for five split scenarios were obtained. The graph in Figure 15 is a result of this analysis.

According to Zibulo’s KPIs, the maximum average away time that should be obtained at any particular time to reach the set production targets is 75 seconds. This KPI is indicated by the horizontal line on the graph. When there are one to three splits between the FB and the CM cut position, the average away times are below the KPI. During the transition from three to four splits, however, the KPI is reached and exceeded before the 4th split is entirely open. Thereafter, the average away times remain higher than the KPI. Figure 15 therefore implies that if the average away times are to be kept below the set KPI, then a belt extension is to be carried out before the 4th split is open. This should be done just after the 3rd split is fully open. At this point, the belt will be extended over two splits as illustrated in Figure 16 (by the red dashed line) such that after the belt extension, one split will be open, allowing for the effective use of only

Figure 16—Illustration of a belt extension with respect to the splits

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Figure 15—Average away times vs. the number of open splits

Optimization of shuttle car utilization at an underground coal mine techniques to be employed are still being determined. Unlike the more established Anglo American Thermal Coal underground operations such Goedehoop Colliery, Zibulo still outsources some of the crucial tasks that are directly linked to operational efficiencies. Cable repairs and maintenance are carried out by the contracting company, Lectropower (Smit, 2014). The company runs an underground mine workshop (referred to as the cable shop) to which cables are delivered for repairs.

The cable management system, according to the author and a student who conducted a project on cable management, is somewhat ineffective. The basic ideal handling of damaged cables at Zibulo Colliery is as follows: (i) When a cable is damaged in a section, it is reported to the cable shop (ii) The cable is then manually loaded onto a load haul dumper (LHD) at the section; the LHD transports the damaged cable from the section to the cable shop (iii) At the cable shop, the cable is offloaded. A spare cable is then manually loaded onto the LHD, which returns to the section (iv) Once the cable has been repaired and tested, the section is notified and the cable is then kept in the spare cable zone (Horstmann, 2014).

Some shortfalls were identified. The following are some of the findings that compromise the ideal procedure together with the operational system. ® Damaged cables are not always handled correctly; this consequently often leads to them not being repairable (Horstmann, 2014) ® No particular operator in the section is responsible for the transportation of damaged cables; this means that in the case of a damaged cable, any one of the section’s operators (one of the two CM operators, one of the three SC operators, or any of the two roofbolt operators) is pulled out and assigned to the task of delivering the damaged cable to the cable shop. Normal operation is therefore disrupted ® The LHDs are shared between sections. No section has its own LHD. A section with a damaged cable usually has to wait for a relatively long period for the next section to deliver the LHD. This implies that valuable production time is not used effectively while the section waits on an LHD to deliver the damaged cable to the cable shop ® Cables are manually handled; the removal and installation of a new cable can take up to an hour

® In most cases, there are no spare cables in the section and so a section had to wait for a spare from the cable shop ® Old cables that can still be used are left in the old working sections during section moves (Horstmann, 2014) ® The cable shop floor space is insufficient; there is no space for the CM cables. These cables are then kept outside the shop, thus leaving them vulnerable to damage (Horstmann, 2014).

It is suggested that Zibulo investigates how the more established Anglo American Thermal Coal underground operations manage their cables and implement cable management initiatives.

Anchoring Owing to continuous repairing of cables, cables lengths are usually shortened. From a maximum cable length of 230 m, a maximum of only 200–210 m remains for SC tramming from the anchor point at the FB to the loading zone and back.

Therefore SCs may not be able to reach loading zones when there are three or more splits are between the FB and the cuts.

According to Smit (2014), SC anchoring requires a great deal of experience to redesign. In particular, due to shorter cable lengths (which consequently lead to cars not being able to reach certain cuts), the location of anchor points in the section can be altered so as to enable effective tramming even for the cars with relatively short cable lengths. All the Anglo American Thermal Coal underground operations employ a similar anchoring method to that used at Zibulo Colliery.

Figure 17 illustrates the possible alternative anchoring points that can be explored.

If car Z was anchored on the original position (on the LHS of the FB), it may not be able to reach the 2nd cut from the anchor point. To avoid this situation the car can be anchored a road away from car Y, on the RHS of the FB. This would significantly reduce the SC’s tramming route distance to the 2nd cut. The reduction in the tramming distance implies a reduction in the shuttle car average away times and thus increased productivity rates.

SC Z can also be anchored midway between the FB and cut 2. In this case the SC uses an effective cable length that is half the cable length required to tram the car to the 2nd cut if it was anchored where car Y is anchored, for example. Smit (2014) suggests that the major concerns with the different anchoring systems are:

Figure 17—Alternatives to the traditional shuttle car anchor points

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Optimization of shuttle car utilization at an underground coal mine ® The running over off a car cable by the next car to pass that point ® The increase in change-out points ® Increase in shuttle car away times.

It has been established that change-out points should be kept as close as practically and safely possible to the loading zone; they should also be kept as close as possible to each other. At a change-out point, one SC has to wait on the next to make way for it to tram to the loading zone or the FB. Moving an anchor point midway between the loading point and FB not only reduces the cable length requirements; it introduces more change-out points in the section as illustrated in Figure 17. This may lead to increased away times as cars are to wait on each other at more points.

However, a constant number of SC anchor points can be maintained if all are anchored midway of their respective tramming routes.

As a SC moves away from the anchor point, the cable and thus the anchor point are under tension. This is illustrated in Figure 18. This tension leads to pillar damage as illustrated.

Pillar damage may extend to the anchor point and thus weaken the anchoring to the extent that the anchor is ultimately pulled off from the pillar. This becomes a safety hazard; it can also lead to cable damage.

To reduce anchor tension that is caused by the cable tension as the shuttle car moves towards the CM, a ‘spring effect’ can be introduced. A used tyre, for example, can be used to connect the anchor point and cable as illustrated in Figure 19. The tension in the cable will be absorbed by the tyre and thus the anchor point will not be greatly affected. A spring can also be used as an alternative.

CM. However, due to the Zibulo underground operation standard, a maximum of three splits should be open before a belt extension. This is also due to cable length restrictions; shuttle cars may not be able to reach some cuts when four splits are open.

When only two SCs are used at any given time, it is effective to use the centre and either the right or left car, depending on whether the cut being excavated is on the right or left of the BR. Therefore, the centre car should always be in operation, even when only one car is being used. However, the use of at least two cars at any given time should be maintained.

The procedure for handling damaged cables should be revisited so as to improve cable management. Other cable anchoring options can be explored. The number of changeout points can be maintained if all three SCs are anchored midway between the FB and CM on their respective tramming routes. Different anchoring configurations can reduce SC cable restrictions to enable the cars to reach certain cuts. The average labour complement of 8.1 is sufficient to run a section; however, if production targets are to be met then absenteeism will have to be managed.

Labour management An average of 77.94% of the actual labour complement was utilized per section from April 2013 to December 2013. From a full complement average of 10.4, a labour complement of only 8.1 was achieved. Even though the full complement was reduced to 10 from 11 in August 2013, the labour complement target is still not being reached. A typical shift per section comprises the following: ® ® ® ® ® ®

2 CM operators 2 roofbolt operators 3 SC operators 1 miner (shift supervisor) 1 electrician 1 fitter.

Figure 18—Cable at anchor point under tension

Absenteeism is attributed to sick leave, absence without pay (AWOP), and planned absence. For a section to be barely productive, eight people (two SC operators less) are required.

However, production targets may not be reached because only one SC will be operated. Thus, absenteeism is to be managed and provisioned for. Zibulo Colliery is currently running a project on discipline enforcement by frontline supervisors.

Conclusion

Figure 19—Anchoring through a tyre to introduce a spring effect to the anchor point and reduce tension

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When there is only one split between the FB and CM, the use of a third SC has no effect on the production already obtained by using two SCs. Three SCs are collectively effective only when there is more than one open split between the FB and

Optimization of shuttle car utilization at an underground coal mine Recommendations It is recommended that when only one split is open, two SCs should be used. The third can be scheduled for maintenance.

When two splits are open, all three cars should be used.

However, in cases where this is not possible, a minimum of two SCs should be used at any given time in the section. The belt should be extended after three splits are open. When cutting on the right-hand side of the BR and using only two SCs, use the right-hand and centre cars. When cutting on the left-hand side of the BR and running only two cars, use the left and centre cars. For safety reasons, pull the change-out point back to avoid shuttle cars turning onto CM operators.

Especially in cases such as a left-hand side car not being able to reach the RHS extremities of the panel, other anchoring configurations should be employed. The car should be anchored midway between the FB and CM cutting point; this will reduce the cable length initially required for the car to reach such extremities.

Designated equipment and operators, particularly for cable transportation, are to be employed so as to avoid having operators being pulled out from a section to transport a damaged cable. It should be ensured that each car has three cables; one in use, one spare in the section, and one at the cable shop. Zibulo should continue with discipline

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enforcement f by ffrontline supervisors. Consequence management should be timely, definite, and consistent.

References ANGLO AMERICAN. 2007. Anglo American announces approval of the Zondagsfontein coal project. Press release: http://www.angloamerican.com/media/releases/2007pr/2007-12-07.aspx [Accessed 20 March 2014].

ANGLO AMERICAN. 2013. Zibulo Colliery: Application for Capital Funding For the Purchase of Three New Replacement Shuttle Cars. Internal Report, Anglo American Thermal Coal Operations.

CALLOW, D.J. 2006. The impact of mining conditions on mechanized mining efficiency. Proceedings of the II Underground Operators Conference, Nacrec, Johannesburg, South Africa, 11-12 September 2006. Southern African Institute of Mining and Metallurgy, Johannesburg. pp. 821â&#x20AC;&#x201C;830.

DE LANGE, M.J. 1988. Underground Transport on Large South African Coal Mines. PhD thesis, University of the Witwatersrand, Johannesburg, South Africa.

OLIVIER, J. 2014. Techincal manager, Zibulo Colliery, Mpumalanga. Personal communication.

HORSTMANN, E. 2014. Cable management at Zibulo Colliery. Undergraduate report. University of Pretoria.

SMIT, R. 2014. Section foreman, Zibulo Colliery, Mpumalanga. Personal communication.

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http://dx.doi.org/10.17159/2411-9717/2015/v115n4a5 ISSN:2411-9717/2015/v115/n4/a5

Explosives utilization at a Witwatersrand gold mine

by M. Gaula* Paper written on project work carried out in partial fulfilment of BSc. (Mining Engineering)

Gold bearing deposits of the Witwatersrand basin are generally less than 2m thick and require conventional narrow-reef mining methods for extraction and employ explosives as a means of rock breaking. Optimal utilization of explosives is dependent on the overall design of the blast. The under-utilization of explosives arises when shot-holes are drilled inconsistently, overcharged, and when tamping is absent. This can be rectified by emphasizing the importance of good drilling practices as part of induction programmes and refresher courses. The project was aimed at determining whether or not explosives are being optimally utilized at project site. This was investigated through a study of the properties of explosives, mine standards, and recommendations for usage. Underground observations were made to determine whether or not mine standards were being adhered to. Historic data was obtained to establish the historic relationship existing between the quantity of explosives used (kg) and the production output (m2). This was then compared to the quantity of explosives the mine expects to use per unit of production. The results obtained were analysed to determine the presence and extent of over- or under-utilization. It was found that explosives are being under-utilized at the mine. More explosives are ordered than expected per unit of production. The explosives’ properties are not thoroughly exploited during blasting, thereby requiring the use of more explosives than prescribed.

Keywords blasting practices, explosives utilization, blast design.

Introduction The South African gold mining industry is based predominantly in the Witwatersrand Basin. The gold reefs found in this basin are generally less than 2 m thick and extend to depths in excess of 3 km below surface, with approximate dips ranging between 20 and 25° from surface (MRM, 2012, p.20). Mines that extract deposits of this nature are narrow-reef mines. The project site is one of these mines, and conventional drill-and-blast mining methods are employed. Hand-held pneumatic rock drills are used for face drilling, explosives are used to fragment the rock, and electricpowered scraper winch systems clean the working areas by removing broken rock from the face and tipping it to the orepass system through a system of in-stope tipping points.

The nature of the orebody and mining environment necessitates the use of explosives as a rock-breaking mechanism, thus making explosives an integral part of the mining cycle.

Without them, production cannot take place.

Explosives utilization is the usage of The Journal of The Southern African Institute of Mining and Metallurgy

Objectives The project is aimed at investigating the types of explosives in use at the project site and improving their utilization by at least 10% by determining the following: ® Factors contributing to explosives utilization ® Whether explosives are currently being optimally utilized ® The relationship between explosives and production ® Mine standard pertaining to explosives utilization ® Possible causes and consequences of over- or under-utilization of explosives ® How explosives utilization can be improved by at least 10%.

Explosives consumption The mine has an expected broken rock output per unit of explosives used (de Sousa, 2013).

The ratio of explosives used to production (centares) is obtained empirically using the following parameters: ® Length of drill steel: 1.2m ® Length of drill steel chuck: 0.3 m ® Length of hole: 0.9 m ® Drilling density: 4 holes per m2 ® Shock tubes: 4 tubes per m2 ® Panel length: 30 m ® Panel width: 1 m

* University of the Witwatersrand, Johannesburg.. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Mar. 2015 VOLUME 115

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explosives in a manner that yields the desired results and that exploits every aspect of their ability to break rock. In order to optimize the use of explosives, a thorough understanding of their properties, characteristics, rock-breaking mechanisms, and application is necessary. An understanding of the basic operational functions of explosives will encourage the implementation of techniques that lead to optimal utilization of explosives.

Explosives utilization at a Witwatersrand gold mine ® Burden spacing: 0.6 m.

One case of explosives (25 kg) breaks rock over a span of 10 m2. A 30 m panel would therefore require 75 kg of explosives and 120 shock tubes.

The quantity of explosives required per panel in kilograms would then be obtained from the sum of the explosives used in shot-holes and those used in preconditioning holes: Shot-holes: 1 case = 10 m2 3 cases = 3 m2 Therefore 75 kg explosives would be required for 30 m2 = 2.5 kg/m2 Preconditioning holes: Nine preconditioning holes are expected and there are three cartridges per hole. A 25 kg box of explosives contains 100 cartridges, each with an approximate mass of 0.25 kg.

The mass of explosives in preconditioning holes for the entire panel is

The panel has on average 66 shot-holes and 7 preconditioned blast-holes. The total mass of explosives contained in the panel is obtained as follows.

For the shot-holes:

[2]

where l = length of priming cartridge (cm) ρ = density of cartridge (g/cm3) R = effective radius of shot-hole (cm).

where l = length of column charge cartridge (cm) The total explosives mass required for a panel is the sum of the mass for the shot-holes and of the mass for the preconditioning holes, which equates to 2.725 kg/m2.

It is important to note that this method of calculating the approximate quantities of explosives required to produce the expected output is based on the following assumptions: ® Face preparation, drilling, charging, and timing are per mine standard ® Panel length is maintained at 30 m and stoping width kept constant ® Secondary blasting is neglected ® Blasting of the gullies is not accounted for.

For the preconditioned holes:

Blast design Optimal explosives utilization is dependent on the overall blast design (de Beer, 2013). It is important to ensure that face preparation, drilling, and charging are done correctly.

Face preparation Blast designs may vary for various reasons, one major reason for this being the stope width. The distance between blastholes, also known as the burden spacing (G), G can be obtained as follows:

[1]

where Mc = mass of explosive per metre of blast-hole (kg/m) K= powder factor (kg/m3).

K The explosives in use have a density of 1.15 g/cm3. Using the expression M=ρV V, the mass of explosives contained in a hole and subsequently, a panel can be obtained. Underground observations carried out on the western panel of the 16th level, 31st crosscut are used below to derive the burden spacing of 60 cm as per mine standard.

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Therefore, the total mass of explosives in the panel is

In order to determine the burden spacing, one needs to take into account the powder factor. This is the mass of explosive required to break one cubic metre of rock, and is calculated using the expression:

[3]

The burden spacing is given by

The Journal of The Southern African Institute of Mining and Metallurgy

Explosives utilization at a Witwatersrand gold mine Mc is the mass off explosives contained per blast-hole. The value has been derived by dividing the total mass contained in the shot-holes by the number of shot-holes,

The length of the blast-hole that actually contains the explosive is then found. Since the holes are 1.2 m long and the total length of the combined cartridges is 580 mm, only 1.2 × 0.580 m is fitted with explosives. As a result, the mass of explosive contained per blast-hole becomes

Substituting this and the K value into the burden spacing equation yields:

Figure1 – Mine standards for charging and blasting

This is the maximum burden that the explosives can effectively handle.

The mine has regulatory policies (mine standards) for all activities carried out during the ore extraction process, which should be adhered to at all times. The mine standards for drilling are as follows: ® All drill-holes must be drilled on the position marked on the face and aligned underneath the direction line ® Holes are to be drilled to the full length of the drill steel ® All the holes marked on the face should be drilled ensuring that each hole has the same burden to break ® Holes must be drilled at an angle no less than 75° to the face ® Temporary support is to be installed prior to commencement of drilling.

Charging The mine standards prescribe the following when charging up and blasting (de Beer and Ross, 2012): ® The primer is prepared by inserting the metal end halfway into the cartridge. This should be done in a safe, approved priming bay away from the blast site to minimize the risk of accidental firing, which could be caused by stray currents or electromagnetic radiation ® Blast-holes are to be de-sludged using an aluminium 3way blowpipe and an approved scraper wire. Safety goggles are to be worn at all times when de-sludging blast-holes ® Explosives should then be transported to the working face in elephant bags. The cartridges and accessories should be transported separately in approved containers (elephant bags) ® The primer should be inserted into the hole first and pushed to the bottom of the hole using a square-ended charging stick ® The column charge is then inserted into the blast-hole.

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®

®

® ®

Results The results presented include historic results obtained from the explosives supervisor and observations recorded underground during the project. The expected explosives utilization is calculated based on the ratio used by the mine – 2.725 kg per m2. Ordered explosives are calculated based on order and delivery forms obtained from the mine, and the ratio obtained by dividing the mass of explosives used by the production throughput for the month.

Historic data The data here enables a direct comparison to be made between the planned and actual explosives consumption, based on the planned production output (obtained from the mineral resource management (MRM) department) and the actual production output (obtained from the production personnel at the shaft) for the period from September to December 2013. The graphs were constructed by comparison of the total planned and actual production in relation to the explosives quantities used.

Underground observations Observations were made in two panels, on levels 18 and 16, to gain an understanding of the quantity of explosives used per blast and to determine whether blasting was conducted as per VOLUME 115

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Proper coupling should be ensured by pushing the column charge as far into the hole as is possible without damaging it The remainder of the hole should be tamped to contain gases inside the hole using clay tamping provided by the mine Shock tubes should then be carefully connected to each other. The connector blocks should be more than 10 cm apart Excessive slack between the shock tubes should be avoided in order to prevent whiplash and damage Lastly, the shock tube starter is connected to the charged face, and this is connected to the central blasting system, which is controlled from the control room on surface.

Explosives utilization at a Witwatersrand gold mine Table I

Stoping production results for September 2013 Panel

Total production (m2)

Miner

Explosives expected (kg)

Explosives ordered (kg)

Shock tubes expected

Shock tubes ordered

127 190 102 145 203 99 37 0 89 86

A B C C D E E F G G

346 518 278 395 553 270 101 0 243 234

525 300 500 200 1150 750 0 275 500 250

508 760 408 580 812 396 148 0 356 344

300 300 900 0 1010 400 0 300 0 700

V1 V2 V3 V4 V5 V6 V7 V7 V8 V9

Table II

Stoping production results for October 2013 Panel

Miner

Total production (m2)

Explosives expected (kg)

Explosives ordered (kg)

Shock tubes expected

Shock tubes ordered

A B C C E D D E E F G G H

190 128 95 153 0 185 0 93 156 130 202 51 0

517.8 348.8 258.9 416.9 0 504.1 0 253.4 425.1 354.3 550.5 139.0 0.0

150 300 300 600 600 0 600 0 0 275 0 450 125

760 512 380 612 0 740 0 372 624 520 808 204 0

0 100 0 300 0 0 600 0 0 200 0 300 100

V1 V10 V11 V11 V13 V5 V14 V6 V7 V7 V8 V15 V16

Figure 2 – Comparison of expected and actual quantity of explosives ordered for September 2013

Figure 4 – Comparison of expected and actual quantity of explosives ordered for October 2014

Figure 3 – Comparison of expected and actual number of shock tubes ordered for September 2013

Figure 5 –Comparison of expected and actual number of shock tubes ordered for October 2013

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Explosives utilization at a Witwatersrand gold mine Table III

Stoping production results for November 2013 Panel

V1 V17 V7 V11 V11 V5 V18 V7 V19 V8 V15 V16 V20 V10 V21 V14 V22 V23

Miner

Total production (m2)

Explosives expected (kg)

Explosives ordered (kg)

Shock tubes expected

Shock tubes ordered

A A B C C D F F

0 0 97 148 228 243 0 107 6 164 77 0 40 102 0 0 0 0

0 0 264.33 403.3 621.3 662.18 0 291.58 16.35 446.9 209.83 0 109 277.95 0 0 0 0

550 0 275 200 750 250 0 250 0 0 300 125 125 50 125 800 600 50

0 0 388 592 912 972 0 428 24 656 308 0 160 408 0 0 0 0

300 0 0 200 300 200 0 200 0 0 100 0 100 100 0 300 0 0

G G H H B I D J B

Figure 6 â&#x20AC;&#x201C; Comparison of expected and actual quantity of explosives ordered for November 2013

Figure 7 â&#x20AC;&#x201C; Comparison of expected and actual quantity of shock tubes ordered for November 2013

Table IV

Stoping production results for December 2013

V1 V17 V7 V3 V4 V5 V6 V18 V7 V24 V8 V25 V15 V16 V20 V22 V11 V21 V11 V14 V22

Miner

Total production (m2)

Explosives expected (kg)

Explosives ordered (kg)

Shock tubes expected

Shock tubes ordered

A A B C C D B F F D G G G H H J C I C D D

86 39 105 68 152

235 106 286 185 414 0 491 87 267 226 0 0 0 0 205 747 0 0 0 0 0

0 0 225 0 0 100 0 0 125

344 156 420 272 608 0 720 128 392 332 0 0 0 0 300 1096 0 0 0 0 0

200 0 0 0 0 0 0 0 100

180 32 98 83 0 0 0 0 75 274 0 0 0 0 0

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Panel

Explosives utilization at a Witwatersrand gold mine

Figure 8 – Comparison of expected and actual quantity of explosives ordered for December 2013

Figure 9 – Comparison of expected and actual number of shock tubes ordered for December 2013

Figure 10 – Comparison of expected and actual explosives used per centare

Figure 11 – Comparison of expected and actual number of shock tubes used per centare

Table V

V2 breast panel Panel characteristics Panel length (m) Stoping width (m) Number of marked holes Number of preconditioned holes Average burden spacing (cm) Number of cartridges used Number of shock tubes used Advance (m)

Week 1

Week 2

Week 3

Average

20 1.2 77 7 58 200 100 0.8

20 1.1 76 7 60 200 100 0.8

19 1.1 70 6 60 180 100 0.8

19.7 1.1 74 7 57 193 100 0.8

Week 4

Week 5

Week 6

Average

11 1.2 42 3 58 100 50 0.8

11 1.2 45 3 60 100 50 0.76

15 1.2 40 3 55 100 50 0.8

13 1.2 43 3 58 100 50 0.79

Table VI

V20 wide raise Panel characteristics Panel length (m) Stoping width (m) Number of marked holes Number of preconditioned holes Average burden spacing (cm) Number of cartridges used Number of shock tubes used Advance (m)

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Explosives utilization at a Witwatersrand gold mine

Analysis of results In September 2013, the expected explosives utilization was exceeded by 51.3%. This was calculated by direct comparison of the explosives ordered and the production throughput (Table I). The general trend for the month was that more explosives were ordered than expected. Eleven cases of explosives and 300 shock tubes were ordered for panel V7, yet there was no production from that panel. The mine records this as explosives unaccounted for (wasted). Upon investigation, it was found that the 16th level is seismically active with bad ground conditions. This particular panel had been badly affected by a seismic event and had been closed, and the miner was assisting in panels V2 and V7, since all three panels are on the same working level. Panel V2 ordered only 58% of expected explosives, and V7 ordered no explosives. Possibly, the first miner was placing explosives orders for the two panels he was assisting in. An explosives order may only be placed by a miner for a workplace officially assigned to him (de Sousa, 2013). Therefore, 11 cases and 300 shock tubes can be accounted for. The remainder of the panels ordered more explosives than expected, and the possible reasons for this are discussed in detail later.

The results obtained for October indicate that more explosives were ordered than expected. There were again panels that received explosives yet showed no production. In this case, V5 and V14 were under the administration of the same miner who received 24 cases of explosives for one panel that were actually intended for another panel. The same applies to panels V13 and V6.

During the month of November, six panels ordered explosives with no production throughput confirming where they have been used. No relationship can be established between panels that ordered explosives without producing and those that produced without ordering explosives. A total of 90 cases of explosives and 600 shock tubes were ordered and these remain unaccounted for. Nothing can be said about their utilization and these explosives can be concluded to have been wasted. December shows the same trend- explosives were ordered yet nothing produced.

Occurrences of November and December are, for the purposes of this report, extreme cases that have required extensive research and enquiries about exactly what happened during that period. The remainder of the cases are those where more explosives were ordered than were expected by the mine.

The ratio of explosives (kg) to production output (m2) expected by the mine is 2.725:1, and 4:1 for shock tubes.

Figures 10 and 11 indicate the performance of the mine in relation to the expected figures. Variations in the ratios are evident, indicating cases of both over- and under-utilization of The Journal of The Southern African Institute of Mining and Metallurgy

explosives. Over-utilization occurred when fewer f explosives were used than expected, and under-utilization when more explosives were used than planned. The contributing factors to both over- and under-utilization of explosives, based on mine standards and underground observations, are discussed in detail below.

Inconsistent blast-hole length and drilling angle Underground observations made revealed that at times, the blast-holes are drilled to a shorter length than specified in the mine standard. The impact of shorter blast-holes is demonstrated using the following simple example.

The ideal case (according to mine standard), assuming a 30 m long panel with a 1m stoping width, is as follows: ® Blast-hole length: 0.9 m ® Advance per blast: approx. 0.8 m ® Explosives used per blast: 29 307.21 g ® Advance over 20 blasts: 16 m.

The effect of short blast-holes can be seen from the following calculation: ® Blast-hole length: 0.85 m ® Advance per blast: 0.75 m ® Explosives used per blast: 29 307.21g ® Advance over 20 blasts: 15 m When blast-holes are drilled shorter than prescribed by the mine standards due to incorrect drilling angles, the advance is reduced although the same quantity of explosives is used as for the full-length blast-holes. This results in under-utilization of explosives because the full potential of the explosives is not used. The calculation above (case 2) is exaggerated slightly because it assumes that all holes in the panel are drilled at 0.85 m length. However, this calculation demonstrates the effect of shorter blast-holes on the utilization of explosives. In addition, if blast-holes are drilled to insufficient lengths, 4.8 cm of face advance is lost per blast (de Beer, 2013).

Incorrect burden spacing For every 10 cm increase in burden spacing, 10% face advance is lost per blast (de Beer 2013). A burden spacing of 60 cm ensures optimal fragmentation, due to the interaction between adjacent charges.

When the burden spacing is increased, the explosive energy needs to travel further than 0.3 m to effectively break rock from the adjacent blast-hole. Thus the explosive energy is depleted before optimum fragmentation is achieved. This results in poor hangingwall and footwall conditions and an uneven face shape, as well as over-utilization of explosives.

If the burden spacing is reduced, the explosive energy released is more concentrated, leading to finer fragmentation off the rock mass, but also to overbreak of the hangingwall. Any deviation from the prescribed burden spacing results in approximately 10% overutilization of explosives, an uneven face shape, and poor fragmentation.

Overcharging Much of the explosives energy concentrated in the blast-hole is not evenly distributed but is concentrated within the confines of the surrounding rock mass. As has been observed underground, there is a misconception that overcharging is beneficial to the advance achieved. However, when more cartridges are placed in a blast-hole than the quantity required, more energy is released into the blast-hole. This energy, if VOLUME 115

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mine standard, as well as to investigate the effects ff off not following the mine standard. Underground observations were limited to two panels because monitoring of the input and output parameters and subsequent analysis was to be done over a series of blasts to increase the accuracy of the results.

Observations were recorded for each shift spent in the respective working places. The weekly averages were then calculated and from these Tables V and VI were compiled. The number of marked holes is inclusive of the holes marked to blast the gullies in both cases, but excludes the contribution of secondary blasting.

Explosives utilization at a Witwatersrand gold mine tamping is sufficient, ff causes both overbreak and fine f fragmenf tation, as were as over-utilization of explosives.

Drill bit deterioration The drill bits used in the working places are 34 mm in diameter. According to Jijingubo (2013) deterioration due to wear and tear results in the gradual reduction of the drill bit diameter, thus causing a reduction in the diameter of the blasthole. Jijingubo suggested that this reduces free movement of the cartridge inside the blast-hole, thus rendering explosives less effective than they would be when using fairly new drill bits.

Poor or no tamping The importance of tamping should not be underestimated.

Underground observations showed that adequate tamping of blast-holes is often neglected when charging up, especially close to the end of the shift. Figure 12 illustrates the importance of tamping.

Explosive energy released into the blast-hole uses two primary mechanisms for rock fragmentation: shock and heave.

For effective fragmentation, the explosive energy should be contained in the blast-hole long enough to cause expansion of the cracks induced by the shock mechanism. Tamping aids in this regard by enabling the explosive itself and the energy it releases to remain in the blast-hole and cause expansion as the gaseous products from detonation penetrate the induced cracks. The absence of tamping or even poor quality installation of tamping allows the gas to escape and hence the energy is released into the surrounding environment. This sometimes causes damage to permanent support elements and overbreak, because the energy is not fully released into the blast-hole but is allowed to escape to other areas where it is not desired.

Figure 13 illustrates the effect of tamping on face advance.

Because the absence of tamping allows gases to escape, the end of the blast-hole is often not blasted, leaving sockets behind and subsequently reducing the advance achieved per blast. For 0.9 m holes, no tamping results in 12 cm loss per blast (de Beer, 2013).

Unused cartridges and shock tubes remaining at the face The mine standards require that unused explosives and accessories be returned to the explosives box and locked away.

Figure 12 – The effect of tamping on explosives effectiveness (de Beer, 2013)

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Strict explosives control policies are employed at the mine – all explosives should be accounted for. The miners keep a record of the quantity of explosives and accessories in storage, and upon receipt of a new batch the quantities are adjusted accordingly. A record of explosives used is to be kept as well.

Underground observations proved non-compliance to this requirement, since in both panels observed, no unused explosives were returned to the explosives boxes, and it was assumed that all explosives and accessories taken into the face were used.

Blasting of gullies and secondary blasting unaccounted for The mine standards require that gullies be blasted such that they lead the face. This is to ensure that the ore blasted has a free face to break into. The centre gully should always lead the face, while following the survey line pegs (Figure 14). In practice, the quantity of explosives used to blast an entire panel includes the explosives used to blast the gully, as well as the face. However, the means of determining the quantity of explosives required per square metre does not distinguish explosives used for blasting gullies. Thus the results overestimate the utilization of explosives to blast the face, whereas some of these were used to keep the gully ahead of the face.

Blasting of the gullies is such an important aspect of production that this usage should be allocated an explosives consumption factor.

The blasting of gullies in a 20 m panel entails five blastholes and would consume approximately 12–15 cartridges, 5–7 shock tubes, and detonating cords. This may appear insignificant, but it increases the amount of explosives used while not contributing to production. The resulting higherthan-expected explosives utilization factor can be corrected by including the explosives used to blast gullies (centre and strike) in the planning of the quantity of explosives expected to be used in a panel.

Secondary blasting is usually unaccounted for when allocating explosives to working places. This is done when removing obstructions such as large rocks from grizzlies and when blasting bad hangingwall conditions, including brows.

Unlike the case of blasting gullies, secondary blasting is used only irregularly and is considered a result of poor primary blasting. However, it is a contributing factor to the apparent over-utilization of explosives.

Figure 13 – The effect of tamping on face advance (de Beer, 2013)

The Journal of The Southern African Institute of Mining and Metallurgy

Explosives utilization at a Witwatersrand gold mine Recommendations

Limitations of the record-keeping/monitoring system The mine has a record-keeping system in place in which all miners order their explosives and accessories for specific workplaces. As can be seen from the results obtained, some workplaces have placed orders for explosives while there is no production to account for the usage. However, it is common for a miner in charge of multiple panels that are relatively close to each other to order explosives for panel A but use them to blast panel B. Comparisons of the expected and actual quantity of explosives used per panel exaggerate the extent of the problem, since the trade of explosives between panels is not taken into account.

Conclusions Explosives are a vital component of hard-rock mining operations using conventional mining methods for ore extraction. Mine standards are in place to ensure that all activities involved in the production process are carried out in a way that ensures employee safety and maximizes production output. This study indicates that explosives are not being utilized to their full capacity at the mine. The biggest contributor to the apparent under-utilization of explosives is the limitations of the system that tracks the usage of explosives underground.

The system does not allow a miner to order explosives unless they are for a specified panel officially assigned to him.

There are currently no means of determining how much of the ordered explosives is actually used underground and how much is returned to the explosives boxes. Other factors contributing to under-utilization of explosives are directly related to the overall blast design. These include, but are not limited to; overcharging, incorrect drilling lengths and drilling angles, secondary blasting not being accounted for, as well as the somewhat impractical expectation of explosives consumption that the mine currently has. Under-utilization of explosives also leads to poor ground conditions and increased costs because the mine has to purchase more explosives than required yet the production output remains unchanged.

The utilization of explosives can be improved by implementing changes in the explosive ordering process and providing a means of tracking whereby ordered explosives are used. By so doing, no explosives will be unaccounted for and the utilization problems encountered in the stopes can be addressed with a realistic picture of the extent of under-utiliation.

The Journal of The Southern African Institute of Mining and Metallurgy

The explosives usage calculator The explosives usage calculator can be introduced into the system to aid in monitoring of explosives usage underground.

This form (Figure 15) would be made available together with the explosives order form. After blasting, the form should be inserted into the communication book at the end of the shift.

The availability of this information is aimed at encouraging the miner to directly monitor explosives usage and compare it to that which is expected.

Adjustment of consumption parameters Planning for explosives consumption at the mine is somewhat unrealistic. The benchmark of 2.725 kg/m2 is based on a panel length of 30 m and constant stoping width. This is not a true reflection of the mining conditions, since pillar extraction is the predominant mining method and the panel lengths are constantly adjusted owing to ground conditions and intersection of geological structures (Tsibuli, 2013). Instead of a fixed benchmark, the mine can employ a consumption calculation method that allows for flexibility due to changing localized conditions and accounts for the blasting of gullies as well as secondary blasting. This will present a practical model from which consumption parameters can be calculated and reduce the apparent extent of under-utilization, thereby improving utilization in future.

Training Formal training Scientific details of rock-breaking should be included in induction programmes and refresher courses to broaden the knowledge of explosives handling personnel and help them understand the importance of a 60 cm burden spacing. Miners should constantly be reminded that overcharging is in no way beneficial to mining operations. In addition, employees should be informed and constantly reminded about the financial implications of face advance loss per blast and how this affects them.

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Figure 14 – Marking of the centre gully

The results of this project indicate that the current system used by the mine to calculate the amount of explosives that should be used per square metre has the following limitations: ® It is based on a panel length of 30 m, which is not the average panel length for the shaft ® Blasting of gullies and secondary blasting is not accounted for when calculating the expected explosives consumption ® Once explosives are delivered to the miner, no further records are kept of their distribution among the various working places ® Miners are permitted to order explosives only for the panels officially assigned to them. The system assumes no trading of explosives takes place between miners.

These limitations exaggerate the extent of explosives unaccounted for and the extent of under-utilization. In order to improve the utilization of explosives, it is important that explosives are used to obtain the best results and not underestimated. The mine can apply the following measures to improve the utilization of explosives.

Explosives utilization at a Witwatersrand gold mine

Figure 15 – The explosives usage calculator

I f Informal l training i i

A k Acknowledgements l d

Diagrammatic representations in the form of clearly visible laminated posters at waiting places and in the change houses informing employees about the impact of poor drilling practices on the centares they produce monthly and their inability to reach set targets.

Models made of rubber, clay, or any recyclable material displayed at various places in the shaft. These should be designed such they show the goal (reaching the mine call factor) and all the factors that prevent the set targets beting reached, such as incorrect burden, shorter shot-holes, poor tamping, overcharging etc. c These factors could be represented by e.g. parasites feasting on the target – something everyone can relate to and work together against.

Introduction of light, flexible 60 cm long strings made of recyclable materials that can be folded into 10 cm or 5 cm portions. These would be made available to all stoping crews.

The aim here is to involve the crew in adhering to a consistent 60 cm burden spacing, and holding the miner accountable for any inconsistencies, which can then be raised by the crew instead of production supervisors. This is an example of the bottom-up management approach.

EurIng C.R Beaumont: Project Supervisor Mr C.K Tsibuli: Production Supervisor, Project Site Mr D. Setshoantsho: Production Supervisor, Project Site Mr R. de Beer: Blasting Design Engineer, Project Site Mr B. Prout: Lecturer.

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References DE BEER, R. 2013. Blasting Design Engineer, Project Site. Personal communication, December 2013.

DE BEER , R. and ROSS, A. 2012. Project Site, Charge up and Connect Blast. pp.

2–5.

DE SOUSA, J. 2013. Explosives wSupervisor, Project Site. Personal Communication.

DE SOUSA, J. 2013. Explosives Supervisor, Project Site. Underground Visit Project Site Report. p. 3.

PROJECT SITE MINE MINERAL RESOURCE MANAGEMENT TEAM. 2012. Competent Person’s Report on the Project Site. p.20.

JIJINGUBO, X. 2013. Miner, Panel V20, Project Site. Personal communication, December 2013.

TSIBULI, C.K. 2013. Production Supervisor, Project Site. Personal communication, December 2013. N The Journal of The Southern African Institute of Mining and Metallurgy

http://dx.doi.org/10.17159/2411-9717/2015/v115n4a6 ISSN:2411-9717/2015/v115/n4/a6

Critical investigation into the problems surrounding pillar holing operations

by J.P. Labuschagne*, H. Yilmaz†, and L. Mpolokeng‡ Paper written on project work carried out in partial fulfilment of BSc. Eng. (Mining Engineering)

Synopsis An investigation into pillar cutting was carried out at a platinum mine on the western limb of the Bushveld Complex. The focus was on crush pillar design and implementation in order to ultimately improve the compliance percentage for pillar cutting. The major findings from the investigation suggest that the pillar cutting problem lies with the implementation of the design rather than the design itself. Observations of the practical issues underground that prevent good pillar cutting were made. After these issues had been identified, recommendations to rectify these problems and a few other issues identified during the investigation were provided. The recommendations are aimed at improving the pillar cutting compliance and reducing the likelihood of pillar bursts or pillar runs, which will ultimately create a safer mining environment.

Keywords crush pillar, design, practical issues, implementation.

The mine has a history of substandard pillar cutting practice. A few of the surrounding mines that have experienced pillar bursts due to substandard pillar cutting were taken as case studies in order to gauge the effect of noncompliance and identify the possible causes that can lead to a pillar burst.

The objectives of this investigation were to review the current design of crush pillars as practised at the mine, and to compare the current design with an alternative method.

This further entailed the identification of the practical limitations experienced underground during pillar cutting. Finally, the recommendations to rectify the problems identified are provided.

Mine standards and compliance The investigation was undertaken at a platinum mine on the western limb of the Bushveld Complex (BC), just outside the town of Rustenburg in South Africa’s North West Province (Figure 1).

The mine, designated ‘mine X’, exploits the Merensky and UG2 reefs, which are currently the only reefs of economic importance.

The BC consists mainly of alternating layers of norite, pyroxenite, and anorthosite.

The general stratigraphy of the Complex is shown in Figure 2.

The mine employs the conventional narrow-reef breast mining method. The workings are served by footwall waste rock development. The mine infrastructure consists of a main vertical downcast shaft that extends down to 15 level, 598 m below surface, and a decline that extends from 15 level down to 28 level at a depth of 927.1 m below surface. The Merensky Reef workings are accessed from the decline, whilst the UG2 workings are accessed from the vertical shaft. The support method consists of a regional pillar and chain crush pillar combination in order to support the hangingwall. The method of stoping and support is illustrated by Figure 3 (note the positions of the regional and chain crush (yielding) pillars.

The Journal of The Southern African Institute of Mining and Metallurgy

The workings at mine X are divided into two sections, the Merensky section and the UG-2 section. The Merensky section is mined between 598 m and 927.1 m below surface, and the UG-2 section is mined down to 598 m below surface. The two sections have different support standards. Pillar cutting compliance refers to the percentage of crush pillars that are cut to the mine standards.

Merensky section Crush pillars for the Merensky section are designed to be 4 m in length and 2.5 m wide, with a pillar height of approximately 1.1 m.

The layout of a Merensky section panel can be seen in Figure 4.

* University of the Witwatersrand, Johannesburg, South Africa, Anglo American Platinum Bursary Holder/Trainee. † School of Mining Engineering, University of the Witwatersrand, Johannesburg, South Africa. ‡ Anglo Development Centre, HRD Co-ordinator Mining. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper rreceived Jan. 2015 VOLUME 115

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Introduction

Critical investigation into the problems surrounding pillar holding operations The pillar compliance percentages for 3 months were compiled and are shown in Table I.

UG2 section

Figure 1—Location of Rustenburg in the Bushveld Complex (Watson et al., 2007)

The crush pillars for the UG2 section are designed to be 3 m in length, 3 m wide, and approximately 1.1 m in height.

Figure 5 illustrates the layout of a UG2 section panel.

The explanation is the same as for the Merensky section (Figure 4), with the only difference being the crush pillar dimensions.

The compliance percentages for 3 months were compiled and are shown in Table II. From Table I and Table II it can be seen that the compliance percentages are well below the acceptable standard of 80%, and this situation is indeed a cause for concern.

Figure 4—Merensky Reef crush pillar layout (not to scale). (1) Crush pillar, (2) distance to pillar reference line, (3) holing between pillars, (4) pillar reference line, (5) panel face, (6) advanced strike gully, (7, 8) dip and strike directions Figure 2—General stratigraphy of the Bushveld Complex (Jager and Ryder, 1999)

Table I

Compliance percentages for three months for the Merensky section Year

Month

Compliance, %

2013 2013 2013

September November December

63 49 42

Figure 3—Stoping and support method (Watson et al., 2007)

The position of the crush pillar (1) can be seen in relation to the pillar reference line (4), the holing between pillars (3), the panel face (5), the advanced strike gully (ASG) (6), the dip (8) and strike (7) direction. Note that the distance between the pillar reference line and the side of the crush pillar (2) should be 0.5 m.

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Figure 5—UG2 Reef crush pillar (not to scale) The Journal of The Southern African Institute of Mining and Metallurgy

Critical investigation into the problems surrounding pillar holding operations Table II

Compliance percentages for three months for the UG2 section Year

Month

Compliance, %

2013 2013 2013

June July September

61 62 45

other. The rock types were separated along the vertical plane.

The weaker part of the pillar started failing slowly, leaving only the stronger portion of the pillar intact. At that stage, the pillar was positioned well into the back areas as mining advanced, where soft loading conditions dominate. Owing to its reduced dimensions, the pillar was then within the bursting range. It is suspected that the pillar then burst, causing a seismic event.

Design and analysis

Case studies Two case studies were conducted at mines that are in close vicinity to mine X and where pillar burst incidents had occurred. One involved a confirmed pillar burst at mine Y, and the other one involved a suspected pillar burst at mine Z. Both these incidents occurred during the project investigation. These case studies shed light on some important issues pertaining to the pillar cutting practice.

Mine Y confirmed pillar burst The pillar burst at mine Y generated a 2.9 magnitude seismic event. As a result, six employees were injured; fortunately, no fatalities occurred. This pillar burst was the result of an oversized pillar that was left inside the stope as mining progressed. According to the mine standards the pillar should have been roughly 3 m wide by 3 m in length. The actual size of the pillar was 17.5 m in length by 6.5 m in width, as illustrated in Figure 6. The blue blocks indicate the size of the crush pillars that should have been left according to the mine standards. As can be seen from Figure 6, the pillar that burst (hatched in grey) was unacceptably oversize.

According to Watson et al. (2007, 2010), pillars should crush close to the face (preferably within the first 7 m) under stiff loading conditions in order for controlled crushing to occur. Watson et al. (2010) also show that pillar bursts are likely to occur at 10 m to 14 m from the face under soft loading conditions. ‘Stiff’ and ‘soft’ loading conditions are similar to the concept of loading done by the stiff and soft testing machines used in rock mechanics laboratories.

It can therefore be concluded that, referring to Figure 6, the smaller crush pillars that surrounded the oversize pillar all failed progressively under stiff loading conditions as the face advanced. The oversize pillar remained intact without crushing and moved well into the back area of the panel, where high stresses were accumulating in the pillar. The pillar eventually started crushing, and the excess strain energy that was stored in the foundation rocks of the pillar then released and caused violent failure.

Crush pillars are designed in order to prevent a back-break of the hangingwall, while maximizing the percentage extraction.

Crush pillars are generally used where mining takes place between 600 m and 1000m below surface (Jager and Ryder, 1999). The support layout used is a regional pillar and crush pillar combination as shown in Figure 3. This support layout allows for an increased extraction ratio while still ensuring stability of the hangingwall.

The first issue to focus on when designing a crush pillar is the residual strength of the pillar. The residual strength of a crush pillar is the strength that the pillar has after crushing has occurred. The residual strength must be sufficient to prevent a back break in order to be effective. Note that the crush pillars do not support the entire hangingwall strata to the surface. The regional pillars left on each side of the stope are responsible for the hangingwall support to surface. This means that crush pillars have to support only the tensile zone that exists between regional pillars.

According to Watson et al. (2010), in order to prevent a back break the residual strength of a crush pillar should be between 8 MPa and 13 MPa when pillar lines are spaced 30 m apart. For the purposes of this design, crush pillars are designed with a residual strength of 13 MPa. The residual strength of a crush pillar can be determined by using a formula developed by Salamon (Watson et al., 2010): [1]

Mine Z suspected pillar burst

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Not much is known about the seismic event at mine Z, which occurred in December 2013, but it is suspected that the event can be attributed to a pillar burst. It is speculated that a large pillar with a width–to-height ratio greater than 10 was left behind in one of the panels. This was deemed acceptable, since the pillars with width–to-height ratios greater than 10 are known to be virtually indestructible (Ozbay et al., 1995).

The problem suspected here was that the pillar consisted of two different rock types, one of which was weaker than the

Critical investigation into the problems surrounding pillar holding operations where h = pillar height (m) w = pillar width (m) Cb = the cohesion of the crushed rock material (MPa) (Watson et al., 2010, citing Salamon).

The results obtained in this investigation suggest that pillar cutting is an application problem and not a design problem. For this reason, further investigations were carried out into the practical problems surrounding pillar cutting.

The Cb value was taken as 1.6 MPa, and h as 1 m. The pillar width corresponding to a residual strength magnitude of 13 MPa can then be calculated by trial-and-error and interpolation.

The next step is to find the pillar strength using the pillar width, which was obtained using Equation [1]. For design purposes, the pillars are taken to be square. There are two pillar strength formulae that are used in this design. Equation [2] refers to the slender pillar formula as currently used by mine X. Equation [2] is an adjusted version of the 1972 Hedley and Grant formula (Watson, 2010).

Practical problems in pillar cutting

[2] where k = the design rock mass strength (DRMS) (MPa) h = pillar height (m) w = pillar width (m) β = 0.75 α = 0.5 (Watson, 2010).

The other pillar strength formula that is used for comparison is explained by Watson et al. (2010) as follows (Equation [3]).

[3] where h and w are as defined above L = pillar length (m) he = [1 + 0.2692 (w/h)0.08)h.

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Drilling discipline of rock drill operators When the rock drill operators (RDOs) drill the stope face for production purposes, they drill in the direction of the ASG (strike direction). According to mine standards, holings between crush pillars have to be blasted after every 7 m face advance for the Merensky section, and after 6 m face advance for the UG2 section. This is due to the different pillar lengths (4 m in the Merensky section and 3 m in the UG2 section), with the pillar holings in both sections being 3 m wide. The problem occurs when it is time to blast the holings into the siding to create the crush pillars. The blast-holes are marked on the stope face and on the siding where the holing should be blasted. The RDOs then begin by drilling perpendicular holes into the stope face in the direction of the ASG. When it is time to turn 90° towards the siding of the panel in order to drill the blast-holes for the holing, RDOs tend to turn less than 90°. This results in a holing that is not blasted in the correct direction, but which is skewed in the direction of the face advance. Owing to this skewness, the holings tend to be longer than usual, hence the pillar width-to-height ratios are affected. The pillars that result from this poor drilling technique are longer than designed for. This in turn affects the effective pillar widths, and can hence lead to a pillar burst problem if not addressed.

Pillar reference line pegs lag behind panel advance

Note that the 136 denotes a strength factor that should be altered for a different rock type. In the case of this design, 136 will be replaced by the DRMS that was established for the particular rock type.

The next step in the design process is to determine the average pillar stress (APS) for the specific scenario. This can be derived by using the tensile zone thickness that the crush pillars have to support.

The pillar factor of safety can then be determined in order to ensure that the proposed pillar will fail close to the face under stiff loading conditions. The pillar should then still provide sufficient support resistance based on the residual strength of the crushed pillar.

The results that were obtained showed a range of safety factors between 0.62 and 0.38, which ensure crushing under stiff loading conditions. Both the peak pillar strength equations [2] and [3] delivered similar results for the pillar strength. The pillar width–to-height ratio obtained was 2.15, which correlates well with the mine standards, where a width-to-height ratio range of 2.0–3.5 is acceptable.

The results also showed that as depth increases, the tensile zone thickness decreases, which means that larger inpanel pillars are more acceptable at shallower depths than at deeper levels.

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The pillar reference line has to be parallel to the ASG and a distance of 0.5 m from the side of the in-stope crush pillars.

This reference line is painted onto the hangingwall of the stope in line with pegs that are installed by the surveyors.

These pegs are normally installed regularly to ensure that the pillar line can be extended in a straight line and does not deviate from its intended direction. Offsets are taken from the pillar reference line to the siding of the panel, indicating where the pillars have to be situated. When the pegs for the pillar reference line lag the advancing panel, the line extension cannot be marked accurately. Generally this results in the pillar line deviating from its intended course, either towards the siding of the panel or towards the panel itself.

When the reference line deviates towards the siding of the panel, undersize pillars can be expected. Conversely, if the line deviates towards the panel, then oversized pillars can be expected.

Substandard face marking In some situations, the face marking is substandard due to the team leaders marking the face, and not the miner. As a result, the drill-holes are drilled in the wrong position or direction. Ultimately, the poor marking practice also affects the holing size and shape. Oversize or undersize pillars are inevitable, depending on the way the siding is marked. If the The Journal of The Southern African Institute of Mining and Metallurgy

Critical investigation into the problems surrounding pillar holding operations

Rock removal (scraping) difficulties The crush pillars are designed roughly 3 m × 3 m on both the Merensky and UG-2 reefs. This means that three blasts are required for a holing to be completed since the advance per blast is normally 1 m. This relates into a face advance of roughly 3 m during this three-blast period, since the face also advances roughly 1 m per blast. Therefore, in order to remove the broken rock from the blasted holing, the face scraper has to be moved back approximately 5–6 m from the face and rigged suitably at the holing after scraping the face.

This takes extra time and effort, and the workers underground in some cases prefer not to lose this time in order to scrape the small quantity of rock, which eventually builds up in the holing.

Scraping the blasted rock in the holing usually causes a more serious problem – the sticks (elongates) become scraped out during this practice. The line of sticks needs to be constantly advanced as the face advances. The maximum distance that these sticks may be from the face is 4.3 m.

When re-aligning the face scraper to remove the blasted rocks after the second and third blast of the holing, the scraper will have to be moved back. This means that it is highly likely to scrape out the sticks, which then have to be re-installed. In addition to the safety risks, the re-installation process takes time and wastes supplies, and hence costs are also raised.

The solutions applied underground to correct these kinds of problems are often crude in the sense that the solution creates another problem somewhere else. For example, the pillar holings could be blasted at an angle to try and avoid the problems associated with rock removal. This in turn could lead to incorrectly and unevenly sized pillars, which are in danger of bursting.

Conclusions When attempting to design in-stope crush pillars, the determination of the tensile zone thickness becomes important in order to evaluate the demand required from crush pillars. As shown in this paper, the tensile zone thickness decreases with increasing depth. This is the premise upon which crush pillars can be implemented in stopes deeper than 600 m with fairly good results. The most important consideration when designing in-stope crush pillars is the residual strength that is required from the pillar in order to arrest a back break. This residual strength is matched to the required pillar width-to-height ratio. The peak pillar strength can then be computed for the width-to-height ratio required, and thus compared to the average pillar stress to determine whether the factor of safety is adequate. It should be noted that in crush pillar design, the safety factor should be less than 1.0, and optimally around 0.7. The low factor of safety is necessary in order to prevent a pillar burst, and to promote pillar crushing close to the stope face under stiff loading conditions.

Important considerations in crush pillar design were highlighted by the case studies of the pillar bursts at mine Y and mine Z. The problem at mine Y was that the pillars were cut with inconsistent dimensions, and one of the oversized The Journal of The Southern African Institute of Mining and Metallurgy

pillars burst in the back area under softer f loading conditions.

The suspected pillar burst at mine Z showed that particular attention should be paid to avoid leaving pillars in situ that consist of more than one rock type.

The mine standards on crush pillars are found to compare well to the design results achieved in this investigation.

However, the poor pillar cutting track record at mine X would lead to pillar burst problems in the near future in view of the incidents of crush pillar failures at the surrounding mines.

Pillar cutting compliance therefore has to improve. During the investigation it was found that the problems regarding pillar cutting is not due to the design, but rather to implementation.

Recommendations The following recommendations are offered to rectify the key issues identified surrounding pillar cutting that can cause pillar bursts and subsequent seismic events. ® The practical problems that were identified during this investigation have to be addressed to improve pillar cutting practice in order to avoid future crush pillar failures at the mine ® The tensile zone thickness increases as the mining depth decreases. This means that in-stope crush pillars at shallower depths carry higher loads than the deeper ones. The safety factor will therefore decrease as depth decreases if the crush pillar size remains the same. The crush pillar width-to-height ratio could be increased (so as to create a stronger pillar at shallower depths) to maintain a safety factor of about 0.7. It is therefore recommended that the width-to-height ratio of in-stope crush pillars is determined separately for each mining level. Mine management would then need to ensure that shift supervisors and mine overseers are aware of the different sizes of in-stope crush pillars on different levels ® Emphasis should be placed on cutting pillars consistently according to mine standards. Larger pillars should not be left in situ in order to compensate for smaller pillars left previously. The leaving of pillars with inconsistent dimensions increases the likelihood of premature pillar failures. The mine standards should rather be applied, even if the pillars were cut oversize or undersize previously ® Special care has to be taken to ensure that pillars are not cut in a position where they consist of more than one rock type. If this cannot be avoided, additional support such as thin spray-on liner or wire mesh and lacing could be applied to improve the crush pillar’s yielding capability ® Pillars that have been identified as oversize pillars within bursting range should be de-stressed by means of drilling two parallel blast-holes into the pillar and blasting with a low powder factor in order to induce crushing. This preferably has to be done before the pillar moves more than 7 m away from the advancing face, where soft loading conditions will start to have an effect ® Drilling and blasting the holings from the top strike gully of the adjacent panel heaves the rock straight into the top gully of that panel, which solves the rock VOLUME 115

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spacing between consecutive blast-holes is too large, then the pillars will tend to be undersize, and when the spacing is too small, the pillars tend to be oversize.

Critical investigation into the problems surrounding pillar holding operations

®

®

®

®

removal problems identified. f Any rock left f inside the siding can be removed by a hand-shovel more easily down-dip The RDOs should be re-trained in order to emphasize that the pillar holings have to be blasted perpendicular to the face. Also, the RDOs should be made aware the dangers of leaving oversize or undersize pillars. This training should preferably be done by rock engineers Regular checks should be made to see whether the pegs are installed so that the pillar reference lines can be extended according to plan. More surveyors could be appointed to ensure that the pegs of the pillar reference lines do not fall behind the plan. Better communication between miners and the survey department should be encouraged so that pillar reference lines can be kept up to date Disciplinary measures should be taken against repeat offenders. This action can be justified by the panels that actually comply with the mine standards. A new bonus system could also be introduced in order to motivate employees to cut the in-stope crush pillars according to mine standards The substandard face marking problem can be overcome by either appointing more miners to share the work load or by training team leaders for face

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marking. Miners can then focus f on marking the pillar holings since the team leaders can mark the face. This reduces the work load on the miner and hence leaves no excuses for substandard marking of the face or holing ® Using a burn cut blasting pattern with 3 m long holes can be trialled in order to see whether it can solve the rock removal problems. If this succeeds, then a holing can be blasted in a single blast, and no rock removal problems will be experienced.

References JAGER, A.J. and RYDER, J.A. 1999. A Handbook on Rock Engineering Practice for Tabular Hard Rock Mines. Safety in Mines Research Advisory Committee, Johannesburg.

OZBAY, M.U., RYDER, J.A., and JAGER, A.J. 1995. The design of pillar systems as practised in shallow hard-rock tabular mines in South Africa. Journal of the South African Institute of Mining and Metallurgy, vol. 95. pp.7–18.

WATSON, B.P. 2010. Rock behaviour of the Bushveld Merensky Reef and the design of crush pillars. Faculty of Engineering and the Built Environment, University of the Witwatersrand, Johannesburg. pp.201-207 WATSON, B.P., KUIJPERS, J.S., and STACEY, T.R. 2010. Design of Merensky Reef crush pillars. Journal of the Southern African Institute of Mining and Metallurgy, vol. 110. pp. 581–591.

WATSON, B.P., ROBERTS, M.K., NKWANA, M.M., KUIJPERS, J., and VAN ASWEGEN, L. 2007. The stress-strain behaviour of in-stope pillars in the Bushveld platinum deposits in South Africa. Journal of the Southern African Institute of Mining and Metallurgy, vol. 105. pp. 187–194. N

The Journal of The Southern African Institute of Mining and Metallurgy

http://dx.doi.org/10.17159/2411-9717/2015/v115n4a7 ISSN:2411-9717/2015/v115/n4/a7

LHD optimization at an underground chromite mine

by W. Mbhalati* Paper written on project work carried out in partial fulfilment of BSc. (Mining Engineering)

Conzal Mine was not meeting production targets in 2013, and it was established that this was caused by the inability of the load haul dump machines (LHDs) to tram the required tonnages. An investigation of the LHD productivity was therefore conducted to identify the inhibiting factors.

This was accomplished by carrying out a literature review on LHD operations to gain in-depth knowledge and conducting observations in the working environment. The relevant information and data on the LHD type used at Conzal was also acquired.

The major inhibitor was found to be excessively long tramming distances in all the sections of the mine. The one-way tramming distances were all significantly greater than the 90 m set in the mine’s code of practice (COP), with the Main Shafts section having the longest average one-way tramming distance of 260 m. The other inhibitor was LHD utilization, which in 2013 was only 47% against a target of 70%.

Simulation of the LHD operation, taking these two factors into account, showed that production could be increased by more than 100%. As a result, it was recommended that conveyor belts should be extended regularly in order to keep tramming distances within the recommended one-way distance of 90 m. In addition, utilization can be improved by minimizing employee absenteeism as well as by modifying the travelling routes such that LHDs do not encounter unnecessary delays.

Keywords underground transport, tramming, load-haul-dump optimization.

Introduction This report focuses on the factors inhibiting the productivity of load haul dump machines (LHDs) at Conzal Mine from achieving set targets. The aim was to identify the major factors contributing to low productivity and recommend ways of reducing or eliminating them.

Background Conzal Mine1 is located near the town of Steelpoort on the eastern limb of the Bushveld Complex, approximately 350 km north-east of Johannesburg, in Limpopo Province (Figure 1).

The mine exploits the MG1 and MG2 chromite seams of the in the Rustenburg Layered Suite of the Bushveld Complex. Conzal Mine historically consisted of three shafts:

1

The actual mine name is omitted owing to the confidential nature of the information in this investigation.

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Project objectives Production is effectively the heart of all mining operations. It is therefore essential that production targets are met on a daily basis so that the mine remains profitable.

Throughout 2013, Conzal Mine was struggling to meet production targets. It is believed that the main reason for the mine not meeting its production target was due to the LHDs not being able to tram enough ore from the face to the grizzly tips. There was therefore a need to determine main factors that inhibited LHDs at Conzal Mine from meeting their tramming target of 2 200 t/d.

The objectives of the project were to: ® Identify major factors leading to the inability of the LHDs to reach the production target ® Provide solutions and recommendations on how to overcome these inhibiting factors.

Methodology In order to achieve the above objectives, the following methodology was employed:

* University of the Witwatersrand. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Mar. 2015 VOLUME 115

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Synopsis

namely, the North, South, and Broken Hill shafts. The main focus of the project was on the South and Broken Hill shafts, as the North shaft was closed down after reaching the end of its productive life. Both shafts access the underground workings by means of declines, and the mining method is bord and pillar. Ore transportation at Conzal Mine is by means of a batch and a continuous transportation system.

The batch transportation utilizes LHDs, which load the broken ore from the face and tram it to the tip, which then feeds into the continuous transportation system consisting of conveyor belts. The mine makes use of a total of 11 lowprofile LHDs (illustrated in Figure 2), with three of the eleven used as standby machines and for underground construction.

LHD optimization at an underground chromite mine the mine design suits their usage.

The literature review was instrumental in highlighting common problems in the mining industry that were relevant to the LHD problem faced by Conzal Mine. Some of these common factors are discussed below.

Maintenance Maintenance is a very critical aspect of the availability and productivity of LHDs as it can reduce machine-related downtime (Inductive Automation, 2011, p. 7). It is imperative that maintenance is conducted on a regular basis to ensure that LHDs retain their original performance capability and reduce the rate of wear. To mitigate the effects of machine downtime, Conzal makes use of a maintenance structure that can be classified into three categories, consisting of preventative, periodic, and breakdown maintenance.

Figure 1 – Location of Conzal Mine

Availability The availability of a machine influences the ability to tram the required tons from the face to the grizzly tip. Availability is the total up-time of an LHD expressed as a percentage of the total allotted time for the machine. LHD availability decreases with age, therefore the optimization of availability can defer the need to spend capital on new machinery when the current fleet has reached its end of life (Machine Downtime, 2014). From empirical results, the average availability of an efficient LHD system in the mining industry should be around 80% to 85%.

Utilization The utilization of a machine can be defined as the time during which the machine is in motion (transmission hours) expressed as a fraction of the engine hours. Conzal measures utilization in the following manner: Transmission hourss Utiliation= /Engine hours [1] Figure 2 – LHD 307 tramming ore to the grizzly tip

® A visit underground to inspect the operating environment of LHDs ® A literature review of LHD operations ® Interviews with relevant personnel (mine manager, mine overseers, engineer, GIT, foremen, artisans, mine planner, miners and operators) to gain additional information on LHD performance and mining conditions ® Underground observation of LHD operation ® Collection of statistical data such as monthly tonnages, LHD availability and utilization, tramming distances, and breakdowns ® Data analysis and simulations ® Draw conclusions and provide recommendations based on the data analysis and simulation results.

Literature review The invention of LHDs can be said to be a milestone in mining.

The LHD has since become a dominant machine in mechanized mines and plays an important role in overall mine production (Samanta et al., 2004, p. 1). LHDs are very versatile and powerful machines, capable of working in the most hostile of conditions while producing the required tonnages provided that

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Utilization can depend on many factors, such as the transport sytem design, availability of panels to be loaded, and presence of operators, to name a few. The major issue affecting utilization is the transport system design. Poor transport system design during mine planning can result in too few or too may machines in relation to the number of of producing faces, resulting in underutilization of some of the LHDs (Mathiso, 2013).

Tramming distances and cycle time The tramming distance can be defined as the one-way distance to the dumping point (as in the case of Conzal Mine) or as the two-way distance, which includes the distance travelled back to the loading point. Tramming distances have a direct influence on the cycle time of a machine, and long tramming distances reduce the rate of ore delivery (tons per hour) by the LHDs. This has an overall effect on the LHDs’ ability to meet production targets. LHD efficiency generally starts to drop when the effective one-way tramming distances are greater than about 80 m (Leeuw, 2013). Massive production losses and inefficiencies can occur once the tramming distance exceeds this general rule–of-thumb distance.

Observations, results, and analysis Tramming distances, cycle times, and other underground and surface observations were used to further understand working conditions and obtain data for analysis. Further useful The Journal of The Southern African Institute of Mining and Metallurgy

LHD optimization at an underground chromite mine information f like the monthly tonnages, downtime, and availability and utilization data was obtained from the relevant mine departments.

Monthly tonnage The original hoisted ore target at Conzal Mine was 2 000 t/d, which was later increased to 2 200 t/d when the Broken Hill Shaft was re-opened for production in April 2013. The tonnage statistics for 2013 are given in Table I. From Figure 3, it can be seen that Conzal failed to meet target for of 2013. Only during the months of May, June, and August was the mine able to meet its target. As a result, the variance in the tonnage for the entire year amounted to a value equivalent to one month’s production (31 117 t).

It can be clearly observed in Figure 3 that there was a sudden reduction in the production in December 2013. This is because the mine produces for only half the month due to public holidays and the customary Christmas break in South Africa. When considering tonnages in isolation, it is inconclusive whether the main cause of not meeting target is due to LHDs, as the tonnages produced depend on many other factors, such as: ® Poor or substandard mining operations in the form of bad shot-hole drilling and blasting practices ® Availability of blasted panels for loading as a result of poor ground conditions or even a loss of panels

Figure 3 – Monthly tonnages, 2013

Table I

2013 monthly tonnages for Conzal Mine Month January February March April May June July August September October November December Total

Actual

Target

Variance

31 660 32 350 31 970 40 490 42 630 38 150 46 770 39650 41 620 40 010 39 755 17 245 442 300

38 000 40 000 38 000 42 000 42 000 38 000 50 600 46200 44 000 50 600 44 000 19 800 493 200

-6 340 -7 650 -6 030 -1 510 630 150 -3 830 -6 550 -2 380 -10 590 -4 245 -2 555 -31 117

Figure 4 – Three-year availability graph for LHDs

® Stoppages of mining as a result of the Department of Mineral Resources (DMR) issuing Section 54 notices for dangerous or unsafe working conditions, which means the loss of several production days while the concerns are addressed.

Availability

Three-year LHD availability statistics for Conzal Mine, % Month January February March April May June July August September October November December Average

2011

2012

2013

78 76 60 50 53 47 41 63 69 53 83 84 63

64 73 84 84 67 69 69 66 75 79 64 73 72

69 66 76 85 87 88 81 87 86 90 80 83 82

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The LHD availability was calculated based on the effective shift time. At Conzal the effective LHD shift time for the day and night shifts is 7 hours, and 6 hours for the afternoon shift; hence a total of 20 hours per day is available for tramming.

Table II shows the average monthly availability of all LHDs at Conzal from 2011 to 2013.

Figure 4 shows the availability of LHDs from 2011 to 2013. It can be seen that the availability of the machines has increase by 19%, from 63% in 2011 to 82% in 2013. The improvement was a result of the commissioning of a refurbished fleet in 2012 from Sandvik. The previous machines had been in continuous operation since 2004. As a result, older machines spent more time being repaired due to the high rate of component failure, which contributed to the low 63% availability in 2011. Although there has been a substantial increase in availability since then, it is still below the mine target of 85%.

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Table II

LHD optimization at an underground chromite mine Utilization Effectively, only a maximum of eight LHDs are utilized per shift due to the number of available panels for loading. The selection of the eight LHDs designated for full-time operation was based on the age of the fleet. From the fleet of eleven LHDs, seven have been in operation for more than five years and the other four were commissioned in 2012. As a result, three of the older LHDs are used as standby units in case of breakdowns. Table III shows utilization values in 2013.

The utilization of LHDs at the mine is a major concern as it is critically low. During 2013 Conzal Mine achieved an average LHD utilization of 47%, which is far below the mine target of 70%.

In order to obtain a better analysis of utilization, data was acquired from Mbhazo Mine, which utilizes the same mining method and machinery as Conzal. Only the months of April to August were taken into consideration because of the availability of data from Mbhazo. It can be seen from Figure 5 that the average utilization at Conzal was lower than at Mbhazo, despite the former marginally outperforming Mbhazo in July and August. The average utilization for Conzal for the period under consideration was 45%, while the average for Mbhazo was 58%. However a firm conclusion cannot be drawn from this comparison as the data covers only a short period.

The situation could be different if data for a longer period, such as a year or more, could be compared.

Figure 5 â&#x20AC;&#x201C; Conzal and Mbhazo utilization comparison graph

Tramming distances The tramming distances were measured from the plans provided by the surveyors. These were measured according to the sections in each shaft to get a better understanding of the average tramming distances. Table IV indicates the one-way tramming distances for the various sections at Conzal Mine.

Table III

LHD utilization for 2013, % Month January February March April May June July August September October November December Average

Achieved

Benchmark

Variance

62 54 50 43 43 42 50 47 52 43 35 43 47

70 70 70 70 70 70 70 70 70 70 70 70 70

-8 -16 -20 -27 -27 -28 -20 -23 -18 -27 -35 -27 -23

Table IV

Tramming distances of different sections, m Section

Distance

Broken Hill Strike 2 Strike 4 Strike 5 Main Shafts

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Figure 6 â&#x20AC;&#x201C;Tramming distances for the different sections

The average tramming distance for Strike 2 could not be measured due to the unavailability of an updated plan at that time, as well as the mineâ&#x20AC;&#x2122;s SOP, which restricts personnel from walking on designated LHD routes and thus prohibits underground physical measurements.

From the data collected one can clearly see that tramming distances are a serious concern. For LHDs to operate efficiently the one-way tramming distances should not be greater than the COP of 90 m. At Conzal, the average one-way distances for the sections measured were all significantly greater than 90 m.

This resulted in great losses in efficiency and ultimately implied that the mine is not getting its return on investment from these machines because they are unable to tram the required tonnages to the tips. Figure 6 shows that the Main Shafts section had the longest tramming distances, followed by Strike 5 and Broken Hill.

As stated in the literature review, there is a relationship between long tramming distances and the tonnages produced by each section because long distances increase cycle times and consequently reduce tonnage output. This relationship between these three factors can be seen in Table V, which shows the tramming distances, cycle times, and tonnages produced in each section in December 2013.

It should be noted that Strike 2 had four working panels, whereas the other sections had eight panels each. This meant the section was naturally bound to produce less than the other sections.

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LHD optimization at an underground chromite mine Table V

Tramming distances, cycle times and tonnages produced for December 2013 Section

Distance (m)

Cycle time (min)

Tonnages (t)

188 171 214 260

10.7 5.9 9 7.2 12.2

2 788 1 974 5 135 4 277 3 069

Broken Hill Strike 2 Strike 4 Strike 5 Main Shafts

varying from f operator efficiency ff and attitude to badly designed tramming routes.

Lengthy cycle times have a direct impact on the tonnages trammed, since fewer cycles will be completed, resulting in poor production output. From Figure 8, one can see that there is an inverse relationship between cycle time and production output. As the cycle time increases, the tonnage outputs tend to decrease (apart from Strike 2, which had half the number of panels of the other sections).

Table VI –

LHD cycle times at different sections, min

1 2 3 4 5 6 7 8 9 10 Average

Broken Hill

Strike 2

Strike 4

Strike 5

Main Shafts

11 9.5 10 10.5 15 9 8 11 13.5 9 10.7

6 5 6 5.5 5 6.5 6.5 6 6 6.5 5.9

8 9.5 8.5 8 9 9 11 8 11 8 9

7.5 7.5 8 6.5 7 5.5 7 7.5 7 8 7.2

12.5 11.5 10.5 12.5 12 12 12 14 12 13 12.2

Figure 7 – Section LHD cycle time graph

The data in the table shows clearly that tramming distances had a direct impact on the ability to meet production targets. The Main Shafts had the longest tramming distances and cycle times, which resulted in a lower tonnage than the other sections with an equal number of panels.

Cycle time Several cycle times were recorded by means of a stopwatch over three days during the afternoon and night shifts across all the sections. The means of the readings are tabulated in Table VI. One can see the differences in cycle time according to different sections in Figure 7. The Main Shafts section has the longest average cycle times, followed by Broken Hill and then Strike 4. Lengthy cycle times can be attributed to many factors,

Figure 8 –Effect of cycle time on tonnage output

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Figure 9 – Strike 4 footwall conditions

LHD optimization at an underground chromite mine As mentioned, badly designed tramming routes, which could manifest in the form of a rolling footwall, have bearing on cycle time. For example it was observed that Strike 4, despite having the shorted tramming distance, experienced long cycle times because of several issues. Firstly, the footwall was often was rolling and in some parts was unacceptably steep. Secondly, there was standing water on the footwall, which reduced traction. Figure 9 shows the roadway conditions at the time of the investigation at Strike 4.

Downtime In December 2013, there were several issues, such as lack of spares, that led to certain LHDs experiencing lengthy downtimes. Table VII shows the downtime for different LHDs in the month.

From the pie chart (Figure 10), it can be clearly seen that LHDs 304, 306, 701, and 702 had had the longest downtimes in December 2013. It should be noted that machines 304, 306, and 701 were old and had not being refurbished. Consequently these LHDs were used as standby machines, seeing that they were already spending a lot of time under repair.

Figure 11 shows LHD 306 undergoing repairs after breaking its rear axle. Such breakdowns take approximately a whole shift to repair and if the spares are not available, it could take up to two days to get the machine back in service. For this particular machine, the downtime was two full days as the mine had to wait for the replacement part from the supplier.

These downtimes have a huge impact on the tonnages hoisted. A loss of 352 t per operational LHD was experienced for the entire December period (see Appendix A for the calculation of this value). This illustrates the need for downtime to be minimized in order to maximize outputs.

Figure 10 – LHD downtime pie chart

Figure 11 – LHD 306 undergoing maintenance

LHD output simulations LHD simulations using Microsoft Excel® were conducted to quantify possible improvements in production when the tramming distances were reduced to the standard of approximately 90 m and the LHD utilization increased to the benchmark of 70%. A single month (December) was considered and the simulations were done for all the sections on the mine. In order to complete these simulations, a triangular distribution was carried out on factors such as operator efficiency, LHD bucket fill factor, and swell to simulate their random nature during loading and hauling operation. The

Table VII

Total LHD downtimes for December, hours LHD no.

Total available time

Uptime

Downtime

LHD 304 LHD 305 LHD 306 LHD 307 LHD 308 LHD 309 LHD 310 LHD 311 LHD 312 LHD 701 LHD 702 Total

200 200 200 200 200 200 200 200 200 200 200 2200

137.08 188.37 144.1 182.07 167.91 165.25 184.8 191.37 181.53 150.65 134.35 1 827.48

62.92 11.63 55.9 17.93 32.09 34.75 15.2 8.63 18.47 49.35 65.65 372.52

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basic equations that were used to complete the simulation were the cycle time, LHD payload, and LHD output per month. The calculations, as well as input variables, can be found in Appendix B. It should be stressed that the results of the simulations do not consider other factors such as blasting delays and interruption from engineering and other service department. To obtain a realistic potential LHD output subject to these considerations, more than a thousand random iterations were done for the two different scenarios: ® Scenario 1: the utilization was kept at the thenprevailing average value of 47% and the tramming distances were randomized through a triangular distribution based on the ideal one-way distance of 90 m ® Scenario 2: the utilization was set to a target value of 70% while the tramming distances were randomized as in scenario 1.

Broken Hill Broken Hill section achieved only 2 778 t in December 2013.

The results of the simulations, tabulated in Table VIII, indicate that the LHDs at Broken Hill section could achieve up to 7 866 t in the first scenario (approx. 90 m tramming distance and 47% utilization) and up to 12 612 t for the second scenario (approx.

90 m tramming distance and 70% utilization). This means that an increase of up to 179% can be achieved by just reducing the tramming distances to the optimum, and up to 352% by reducing tramming distances and increasing utilization.

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LHD optimization at an underground chromite mine Table VIII

Table XII

Broken Hill simulation results

Main Shafts simulation results

Scenario 1

Scenario 1 2 778

7 766.30

12 611.55 Normal

442.39

726.96

Table IX

Tons achieved Simulated output tons Distribution Standard deviation

Scenario 2 1 974 9 040.09

14 761.99 Normal

502.67

791.87

Table X

Tons achieved Simulated output tons Distribution Standard deviation

Scenario 2 5 135 8 486.97

13 894.52 Normal

482.97

765.78

Table XI

Strike 5 simulation results Scenario 1 Tons achieved Simulated output tons Distribution Standard deviation

Scenario 2 4 277 7 826.62

12 730.73 Normal

384.80

11 554.15 Normal

368.64

582.02

The Main Shafts section performed relatively poorly in that it managed to produce only 3 069 t in December 2013. The simulated output of the LHDs in this section could be increased by up to 131% for the first scenario and 276% for the second scenario. This demonstrates that the extremely long tramming distances in this section constitute a serious bottleneck in the overall attainment of the mine’s production target. Such long distances translate into very low LHD efficiencies.

Conclusions

Strike 4 simulation results Scenario 1

3 069 7 079.44

Main Shafts

Strike 2 simulation results Scenario 1

Tons achieved Simulated output tons Distribution Standard deviation

Scenario 2

678.05

Strike 2 Strike 2 follows a similar trend to the Broken Hill section.

Compared to the actual output of 1 974 t, the simulation results indicated that Strike 2 section’s production could be improved by 358% in scenario 1 and 648% in scenario 2. Table IX shows the summary for this section.

Strike 4 This section managed to achieve 5 135 t for the month of December. Based on the simulations, this section has the potential to improve by 65% and 171% for scenario 1 and 2 respectively. One can now see the effect of the tramming distances, as the increments are not as high as those for Broken Hill or Strike 2, since this section had shorter tramming distances than the others.

Strike 5 This section follows the same trend as the other sections with the simulations indicating that there would be an improvement in the output tonnages when the two scenarios are applied.

The improvements based on Table XI are 83% and 198% for scenarios 1 and 2 respectively.

The Journal of The Southern African Institute of Mining and Metallurgy

LHD transportation needs to be designed in such a way that it is efficient and meets the production requirements of a mine.

Conzal’s production for 2013 was 442 283 t, against a target off 493 200 t.

Based on the analysis of the tonnages, availability and utilization figures, tramming distances, cycle times, breakdowns, and the LHD simulation results, it was concluded that the major factors preventing the LHDs from meeting the required production target were tramming distances and utilization. LHD simulations showed that the production outputs of all the sections could be increased considerably by decreasing tramming distances and increasing LHD utilization.

This is because tramming distances have a direct impact on all the other issues investigated, and if the tramming is reduced to optimal distances, then all the other issues will mostly likely be alleviated. When the tramming distance is reduced: ® The cycle times will be reduced because the LHDs will be travelling a shorter distance at the same speed, which can translate to more ore being trammed in a given period ® Breakdowns are likely to be reduced because the machines will suffer less wear and tear ® The reduction in breakdowns will automatically improve availability, which will in turn further increase the ore tonnage trammed.

Recommendations Based on the analysis of the results and observations, the following recommendations were made to address the factors that resulted in failure to meet production target. ® Tramming distances for all the sections must be reduced to less than 90 m. This can be achieved by extending the belts on all the sections on a regular basis to ensure that tramming distances are within the optimal distance ® The utilization of the LHDs needs to be improved significantly. A low utilization value is unacceptable for efficient tramming. Utilization can be improved in various ways, including reducing absenteeism and VOLUME 115

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Tons achieved Simulated output tons Distribution Standard deviation

Scenario 2

LHD optimization at an underground chromite mine increasing team spirit amongst the employees. The design of the tramming system also needs to be carefully re-evaluated to address any design errors that could lead to low utilization. ® Maintenance needs to be optimized so that LHD downtime can be reduced. This can be achieved by maintaining an adequate stock of spares in the mine store Cycle time needs to be reduced, and this can be done by improving roadways and introducing an improved housekeeping system to ensure that they are maintained in good condition.

Acknowledgements I would like to offer my special thanks to the following persons and organizations for their assistance with this investigation: ® The company referred to as Conzal Mine, for giving me the opportunity to conduct the investigation and fulfil the requirements of the project ® My project supervisors, P. Leeuw, D. Pretorius, and B. Mathiso for their generosity and dedication in assisting me to produce this report ® My mother and father, Mr and Mrs Mbhalati, for their love and constant support in times of difficulty and struggle ® The mining and engineering crews at the Mine for their assistance in acquiring the data and knowledge required for the project.

Appendix B Template of LHD simulation input variables Effective loading hours per day

14

Number of LHDs

2

Availability

83%

Utilization

70%

LHD bucket size (m3)

2.2

LHD average speed (m/s)

2.2

Two-way tramming distance (m)

177

In-situ density

4.7

Assumptions: Operator efficiency

79%

Bucket fill factor

94%

Swell

1.49

Loading and dumping time (s)

100

Additional time

LHD output per month: 1. Cycle time (h) = [2]

2. LHD payload (t) =

∗ Bucket size (m3) ∗ Bucket fill factor

Appendix A Calculations of the tons lost per LHD for the month of December 2013: ® Total available tramming time for the eight full-time operational LHDs Tt = 1600 hours ® Total downtime for the eight LHDs Td = 204.35 hours ® Total average downtime per LHD

60

[3]

3. LHD output (t/month) =

∗ No. of LHDs ∗ LHD payload ∗ Cycle time ∗ Availability ∗ Utilization ∗ Opperator efficiency [4]

References ® Daily tonnage target =

INDUCTIVE AUTOMATION. 2011. White Papers on Automation, Process Control & Instrumentation Topics. http://www.automation.com/pdf_articles/Whitepaper-Reduce-DowntimeRaise-OEE.pdf. [Accessed 21 April 2014].

Effective loading time per day = 20 hours ® Required tons per hour:

LEEUW. P.K.J.. 2013. Mine Transportation. Notes for Mine Transportation course.

University of the Witwatersrand.

MACHINE DOWNTIME. 2014. Machine Downtime. http://machine-downtime.com/ [Accessed 23 April 2014].

MATHISO, B. 2013. Maintenance coordinator [Interview] 9 December 2013.

® Required tons per hour per LHD,

SAMANCOR CHROMe. 2014a. About Us - Company Overview. http://www.samancorcr.com/content.asp?subID=2 [Accessed 15 April 2014].

The total tons lost for December due to average downtime per LHD can be calculated as follows: ® Tons lost per LHD for December = Td-ave * tLHD = 25.54 h * 13.75 t/h = 351.23 t per LHD

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SAMANCOR CHROME. 2014b. Our Business - Operations and Locations. http://www.samancorcr.com/content.asp?subID=8 [Accessed 15 April 2014].

SAMANTA, B., SARKAR, B., and MUKHERJEE, S.K. 2004. Reliability modelling and perfomance analyses of an LHD system in mining. Journal of the South African Institute of Mining and Metallurgy, vol. 104, no. 1. pp. 1–8. N The Journal of The Southern African Institute of Mining and Metallurgy

http://dx.doi.org/10.17159/2411-9717/2015/v115n4a8 ISSN:2411-9717/2015/v115/n4/a8

The viability of using the Witwatersrand gold mine tailings for brickmaking

by M. Malatse* and S. Ndlovu* Paper written on project work carried out in partial fulfilment of BSc. Eng. (Metallurgy and Materials Sciences)

The Witwatersrand Basin is the heart of South Africa’a gold mining industry. The cluster of gold mines located in the Witwatersrand Basin generates a significant amount of mine tailings, which have adverse effects on the environment and ecological systems. In addition, disposal costs are very high. The exponential population growth in the Witwatersrand area has resulted in pressure on the reserves of traditional building materials.

Quarrying for natural construction material is very expensive and damages the landscape. This work therefore examines the use of gold mine tailings in the production of bricks.

Different mixing ratios of gold tailings, cement, and water were used.

The resulting bricks were then cured in three different environments – sun dried, oven dried at 360°C, and cured in water for 24 hours. The bricks were then tested for unconfined compressive strength, water absorption, and weight loss. The results showed that the mixture with more cement than tailings had a compressive strength of approximately 530 kN/m2. It was also found that the best brick curing system was in a water environment.

Bricks made from tailings cost more than conventional bricks because of the higher quantity of cement used, but the manufacturing process consumes less water. Overall, the results indicated that gold mine tailings have a high potential to substitute for the natural materials currently used in brickmaking.

Keywords gold mine tailings, construction materials, brickmaking.

Introduction South Africa is a mineral-rich country with metals such as gold, copper, and platinum group metals being exploited to a significant extent in the country’s mining history. Mining generates large volumes of tailings, with consequent disposal and environmental problems. By far the most gold that has been mined in South Africa (98%) has come from the Witwatersrand goldfields (Messner, 1991).

The gold mines in this area are situated around an ancient sea (over 2700 million years old) where rivers deposited sediments in the form of sand and gravel that became the conglomerate containing the gold (Messner, 1991). The extensive exploitation of the gold resources has led to numerous mine tailings heaps scattered around the Witwatersrand Basin. As long as mining contributes significantly to the economic development of South Africa, generation of these tailings is inevitable.

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* School of Chemical and Metallurgical Engineering, University of the Witwatersrand, Johannesburg, South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Feb. 2015 VOLUME 115

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Synopsis

The major environmental impacts from waste disposal at mine sites can be divided into two categories – the loss of productive land following its conversion to a waste storage area and the introduction of sediment, acidity, and other contaminants into surrounding surface and groundwater (Mining Facts, 2014). The gold mining and processing wastes contain large amounts of sulphide minerals such as pyrite, which generate acid mine drainage (AMD) (Rosner and van Schalkwyk, 2000). South Africa is currently faced with the challenges resulting from AMD and the government and mining companies are under pressure to find viable solutions to this problem. This, coupled with the increasing landfill costs, and stricter implementation and enforcement of environmental legislation, has caused the scientific community to focus on finding innovative methods of utilizing mine tailings. Even though some applications of the generated tailings have been exploited, such as in the building of slimes dams and backfill in underground mines, these uses do not take up more than a fraction of the total amount of tailings in the Witwatersrand region. There is therefore a significant need to developing other long-term, commercially viable uses for mine tailings in order to minimize the disposal costs and the impact on the environment.

According to Statistics South Africa (2013), South Africa has a human population of about 52.98 million. This population is growing, and this consequently results in an increasing demand for housing, which places severe stress on the natural resources used for construction materials. Conventional bricks are produced from clay fired in high-temperature kilns or from ordinary Portland cement (OPC) concrete. Clay, the common material used for

The viability of using the Witwatersrand gold mine tailings for brickmaking brickmaking, is usually mined in quarries. Quarrying operations are energy-intensive, adversely affect the landscape, and generate a high level of waste (Zhang, 2013; Bennet et al., 2013). Furthermore, in many areas of the world, there is already a shortage of natural resource material for the production of the conventional bricks (Zhang, 2013).

To conserve the clay resources and the environment, some countries such as China have started to limit the use of bricks made from clay (Zhang, 2013). Thus the depletion of these natural resources has created a need need for an alternative source of construction materials in order to sustain development.

Extensive research has been conducted on the production of bricks using waste material (Zhang 2013; Saeed and Zhang, 2012). These waste materials include mining waste, construction and demolition waste, wood sawdust, cotton waste, limestone powder, paper production residues, petroleum effluent treatment plant sludge, kraft pulp production residue, cigarette butts, waste tea, rice husk ash, crumb rubber, cement kiln dust, and coal fly ash (Zhang 2013; Bennet et al., 2013; Saeed and Zhang, 2012). The mining and mineral processing waste includes mining overburden, waste rock, mine tailings, slags, granulated blast furnace slag (GGBS), mine water, water treatment sludge, and gaseous waste ( Zhang, 2013; Saeed and Zhang, 2012; Koumal, 1994; Dean et al., 1968; Bennet et al., 2013).

The extensive research on the utilization of waste materials to produce bricks can be divided into three general categories based on the production methods – firing, cementing, and geopolymerization, Production of bricks from waste materials through firing uses waste material(s) to substitute partially or entirely for clay and follows the traditional method of kiln-firing. Chen et al. (2011) studied the feasibility of utilizing haematite tailings and class F fly ash together with clay to produce bricks. Tests were performed to determine the compressive strength, water absorption, and bulk density of brick samples prepared under different conditions.

Bennet et al. (2013) conducted research on the development of geopolymer binder-based bricks using fly ash and bottom ash. During the synthesizing process, siliconaluminium bonds are formed that are chemically and structurally comparable to those binding the natural rocks (Bennet et al., 2013), giving geopolymer binder-based bricks advantages such as rapid strength gain and good durability, especially in acidic environments. Research into geopolymer bricks has also incorporated copper mine tailings and cement

kiln dust (Bennet et al., 2013). In this process, an autoclaved aerated cement (AAC) material is produced (Koumal, 1994).

Ahmari and Zhang (2012) investigated the utilization of copper mine tailings to produce geopolymer bricks by using sodium hydroxide (NaOH) solution as the alkali activator.

They produced cylindrical brick specimens by using different initial water contents, NaOH concentrations, forming pressures, and curing temperatures. Copper mine tailings bricks have been found to have good physical and mechanical properties such as a water absorption of 17.7%, compressive strength of 260 kg/cm2, and density of 1.8 g/cm3 (Be Sharp, 2012).

The method of producing bricks from waste materials through cementing is based on hydration reactions similar to those in OPC to form mainly C–S–H and C–A–S–H phases contributing to strength (Zhang, 2013). The cementing material can be the waste material itself or other added cementing material(s) such as OPC and lime. Again, many researchers have studied the utilization of waste materials to produce bricks based on cementing. The brickmaking process has involved the use of waste and tailings such as those from copper, nickel, gold, aluminium, molybdenum, and zinc processing as additives replacing some of the cement (Jain et al., 1983). Morchhale et al. (2006) studied the production of bricks by mixing copper mine tailings with different amount of OPC and then compressing the mixture in a mould. The results showed that the bricks had a higher compressive strength and lower water absorption when the OPC content increased. Roy et al. (2007) used gold mill tailings mixed with OPC, black cotton soils, and red soils in different proportions to make bricks. The cement-tailings bricks were cured by immersing them in water for different periods of time and their compressive strengths were determined. Bricks with 20% cement and 14 days of curing were found to be suitable. Gold mine tailings have also been used to produce autoclaved calcium silicate bricks (Jain et al., 1983). The bricks are cured under saturated steam and in the process, lime reacts with silica grains to form a cementing material consisting of calcium silicate hydrate. Some mining companies such as Bharat Gold Mines in India have explored the idea of brickmaking using gold ore tailings (Be Sharp, 2012).

Table I shows the chemical composition of some of the waste materials used in bricks as well as the composition of quarry clay material that comprise the conventional feed material (Bennet et al., 2013). The gold mine tailings are from a Chinese mine (Yang et al., 2011).

Table I

Composition of material used in brickmaking (Bennet et al., 2013; Yang et al., 2011) Oxide component

SiO2 Al2O3 Fe2O3 CaO SO3 FeO MgO Na2O K2O

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Fly ash Mass %

GGBS Mass %

Bottom ash Mass %

Clay material Mass %

Gold mine tailings Mass %

53.3 29.5 10.7 7.6 1.8 -

35.47 19.36 33.25 0.8 8.69 -

56.76 21.34 5.98 2.88 0.72 -

61.8 25 8 1.2 0.1 2.76

38.60 7.06 12.76 29.24 3.21 =7.85 -

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The viability of using the Witwatersrand gold mine tailings for brickmaking

Materials and methods The materials used in this test work were gold mine tailings, water, and cement as a binding material. Gold mine tailings were provided by a local gold mining company, AngloGold Ashanti. The Larfarge 42.5 kN cement was provided by the Planning, Infrastructure and Maintenance Department at the University of the Witwatersrand, Johannesburg. The cement was used on the day of delivery and tap water was used in the mixing process.

Characterization of gold mine tailings Representative samples used in all experiments were prepared using a riffle splitter (model 15A, Eriez Magnetics, South Africa). The gold tailings were characterized by investigating the phase mineralogy, particle size, and quantitative chemical analysis. The particle size analysis was done by physically screening the samples using test sieves (Fritsch, Germany) of various screen sizes up to 212 μm. The phase mineralogy analysis was carried out using an X-ray diffractometer (X’Pert, PANalytical, Netherlands) operated with Co-K radiation generated at 40 kV and 50 mA. The chemical analysis was carried out using wavelength dispersive X-ray fluorescence (XRF) spectrometry (Axios, PANalytical, Netherlands) operated with a rhodium tube excitation source.

The brickmaking process Different mixing ratios of tailings, cement, and water were used in the brickmaking process (Table II). From each mixture, a number of bricks were cast and dried.

The three feed material (tailings, cement, and water) were mixed in the appropriate ratios in a commercial mixing The Journal of The Southern African Institute of Mining and Metallurgy

machine. Dry mixing was done first f and then a controlled amount of water was added while continuing to mix thoroughly. The total mixing time was 15 minutes. The mixture was then cast into the brick moulds. The brick moulds were then placed on a vibrating machine for 5 minutes in order to fill the voids in mixture comprehensively and thus prevent the formation of air pockets. The bricks were then labelled and allowed to cure for 24 hours. Three curing methods were used. These included atmospheric drying under the sun, curing in water, and drying in an oven at 360°C. After curing, the bricks were de-moulded using an air compressor, weighed, and tested for compressive strength.

Unconfined compressive strength testing The cast and cured bricks were tested for compressive strength using a Tinus Olsen compressive strength testing machine. In the compressive strength testing process, a force was applied on the brick until the brick failed and the force measured at failure was documented. The compressive strengths obtained were then averaged. The mixture ratio that gave the highest compressive strength was subsequently employed to manufacture bricks for water absorption, weight loss, and leaching rate tests. Unconfined compressive tests were also done on commercial bricks to provide a basis for comparison.

Water absorption rate and weight loss tests Two solutions with different pH values, one acidic and one neutral, were used for these tests. The tailings bricks were first prepared from mixture 7 (Table II) and cured in water for 24 hours. Tests were conducted on four samples in each solution. The bricks were immersed in water baths, one containing water at pH 7 and the other an acidic solution at pH 4. The solid–to-liquid ratio was maintained at 15. The saturated weight of the bricks (Ws) was measured every 24 hours over a 5-day period. After 5 days, the bricks were dried at 110°C for 24 hours and the oven-dried weight (Wd) recorded. The bricks were again tested for compressive strength. The percentage water absorption rate was then calculated as

Water absorption (%) = [(W Ws - Wd]/W Wd × 100 The weight loss tests were done in the neutral environment only (pH 7). The average weight loss was measured after the bricks had been soaked in neutral water for seven days then dried overnight at 110°C.

Table II

Different mixtures used in brickmaking Mixture number 1 2 3 4 5 6 7 8

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Tailings (kg)

Cement (kg)

Water (L)

2 14 9 7 10 12 5 10

1 2 6 8 5 3 10 5

0.6 2.65 3.0 2.5 2.5 2.5 3.3 3

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From the chemical compositions shown in Table I, it can be seen that the waste materials have similar major oxides in their compositions. The compositions are also relatively similar to the typical clay material used in brickmaking. The waste materials all have a predominantly high content of silica, alumina, and haematite (with the exception of the granulated blast furnace slag, GGBS, which contains no haematite), which are important in brickmaking materials.

Considering the source of the gold deposits in the Witwatersrand Basin (river sediments in the form of sand and gravel), it is therefore likely that the tailings from this area will also contain a high level of silica.

The purpose of this work is therefore is to ascertain the technical and economic viability of using the Witwatersrand gold tailings for brickmaking using the cementing method.

The tailings-based bricks will be compared with the commercial bricks available on the market. The evaluation will be based on parameters such as compressive strength, water absorption, and weight loss tests. This work has the potential to unlock large resources of material needed in the construction industry that would help conserve the natural resources commonly used. In addition it would eliminate the land requirements for waste disposal, thus realizing savings on disposal and landfill costs and also lessening environmental damage. But above all, this work has the potential to provide an additional revenue stream for the gold mining sector.

The viability of using the Witwatersrand gold mine tailings for brickmaking Results and discussion

Unconfined f compressive strength off the gold tailings bricks

Particle size distribution Figure 1 shows the particle size distribution of the material used in the brickmaking process. The results are presented in cumulative form, in which the total amount of all sizes retained or passed by a single notional sieve is given for the range of sizes.

The results indicate that most of the particles fell into 90–200 μm range. 80% of the material passed the 200 μm screen aperture while about 12% passed the 90 μm screen.

The particle size range used in standard commercial brickmaking includes coarser sand particles as well as fine particles. The material used in these tests was, in comparison, relatively fine. A cost analysis study done by Roy et al.

(2007) showed that cement-tailings bricks are generally uneconomical compared to the soil-tailings based bricks, therefore future test work will have to consider the addition of coarse particles, possibly from mining overburden.

Mineralogical and chemical analysis Table III shows the mineral phases and the respective quantities present in the sample as determined by XRD and XRF analysis. The table indicates that the mineralogical and chemical composition of the tailings bear close similarities with the composition of the conventional materials used for commercial brickmaking, as well as with the waste materials that have been tested in the past (see Table I). The results indicate that the major oxides in the mine tailings sample are silica, magnesium oxide, alumina, sulphur trioxide, potassium oxide, calcium oxide, and haematite. The other constituents such as uranium oxide are found in trace quantities. Although uranium oxide is present only at 0.0064% its presence is worth noting as uranium is a very radioactive element and therefore can present safety implications.

Unconfined compressive strength The main mechanical property of bricks that is tested for is compressive strength. A good brick should be hard and strong. The compressive strength tests on commercial bricks were undertaken in order to provide a basis for comparison with the gold mine tailings bricks. Table IV shows the results of the compressive strength of the commercial bricks. It was noted during the tests that the more uneven and rough the surface of the brick, the quicker it failed.

The quality and durability of the concrete mix depend not only on the quality and properties of the ingredients, but also on the method of preparation and the curing environment (Ahmad and Saiful Amin, 1998). Proper curing is indispensable in developing optimum properties. Table V shows the compressive strength for the gold tailings based bricks cured in different environments.

The average values shown in Table V are depicted graphically in Figure 2. For mixture 1, high-temperature drying in an oven yielded the highest compressive strength.

For mixture 2, ambient drying conditions resulted in the highest compressive strength, followed by oven drying for mixture 3, curing in water for mixtures 4 and 5, oven drying for mixture 6, and curing in water for mixtures 7 and 8. The overall trend reveals that the majority of the mixtures yielded higher compressive strength when cured in water (50%), followed by oven drying (37.5%), and lastly drying under ambient conditions (12.5%). This can be attributed to the fact that curing the bricks in water contributes to the cementation process and hence increases the strength of the bricks. An

Table III

Major constituents of the gold mine tailings Number 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21

Component

Result (%)

Na20 MgO Al2O3 SiO2 P2O3 SO3 K2O CaO TiO2 Cr2O3 MnO Fe2O3 Co2O3 NiO CuO ZnO As2O3 Pb2O SrO ZrO2 U3O8

0.613 1.79 10.2 77.7 0.085 0.905 1.19 1.93 0.469 0.45 0.0549 4.51 0.0063 0.0177 0.007 0.008 0.01 0.0041 0.0151 0.0312 0.0064

Table IV

Compressive strength of commercial bricks Brick

1 2 3 4 5 6

Force (kN/m2) Flat face 890 920 665 695 690 641

Figure 1—Particle size distribution of the gold mine tailings

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The viability of using the Witwatersrand gold mine tailings for brickmaking Table V

Average compressive strength of bricks cured under different Mixture

1 2 3 4 5 6 7 8

Average compressive strength (kN/m2) Water

Oven

Ambient

141 20 325 440 262 215 530 149

165 25 359 439 261 235 479 98

157 29 318 323 234 230 454 127

absorption rate was slightly higher in the neutral solution than in the acidic solution. The unconfined compressive strengths after water absorption are shown in Table VI. The results show that the bricks soaked in the neutral environment had a higher compressive strength than those soaked in an acidic environment. This can be attributed to the fact that during the water absorption test, the neutral solution acts as a natural curing agent and further strengthens the bricks.

The weight loss over the seven day period was quite negligible at 0.06%. This means that although the bricks show significant water absorption rate, they regain their original weight after drying.

Cost analysis It is important to check if the outcome of the research project is economically viable for it to be beneficial to society. In order to market the bricks, cost comparison with traditional bricks is essential. The following factors were considered.

Gold tailings are available in abundance and are expected to be free of cost Portland cement=R65 per 50 kg bag (OLX, 2014).

Using a base figure, for commercial brickmaking, the masonry cement recipe can be estimated as follows:

8 bags of cement=1000 bricks (Kreh, 2003), or 1 bag of cement=125 bricks.

Figure 2â&#x20AC;&#x201D;Compressive strength of the cement tailings bricks cured in different environments

adequate supply of moisture is necessary to ensure sufficient hydration for reducing the porosity to such a level that the desired strength and durability are attained The results also show that in general, bricks from mixture 7 had a higher compressive strength in all three curing system used. However, the highest overall compressive strength was obtained from mixture 7 that was cured in water. This mixture had a higher amount of cement compared to the tailings (2:1 cement to tailings mass ratio), which resulted in a larger surface area of the tailings being in contact with the cement and hence resulting in a stronger mixture. These results also follow for mixtures 3 and 4. The higher strength is probably due to the superior plasticity and binding properties provided by the higher amount of cement.

It is also known that cement cures well in water (America's Cement Manufacturers, 2014); hence the mixture with the largest quantity of cement cured in water resulted in the highest compressive strength.

For commercial brickmaking, the water addition should be 20 litres per 50 kg of cement (Hydraform, 2014). The price of water for industrial companies according to the City of Johannesburgâ&#x20AC;&#x2122;s Mayoral Committee is R20.96 per kilolitre.

(COJ: Mayoral Committee, 2013).

Figure 3â&#x20AC;&#x201D;Water absorption rate

Water absorption and weight loss tests

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Table VI

Average compressive strength after absorption tests Solution pH 4 pH 7

Compressive strength (kN/m2) 445 476

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Compressive strength and water absorption are two common parameters considered by most building materials researchers as required by various standards. Water absorption will influence the durability and strength of the bricks. Figure 3 shows the water absorption rate.

For both solutions, the absorption was highest on the first day of the test followed by a more constant rate in subsequent days. It can also be seen from Figure 3 that the

The viability of using the Witwatersrand gold mine tailings for brickmaking In this research project, it was found f that the highest strength was obtained in mixture 7, with 5 kg of tailings and 10 kg of cement mixed with 3.3 litres of water, followed by mixture 4 with 7 kg of tailings and 8 kg of cement mixed with 2.5 litres of water. Using the option with the highest strength, it was found that one bag of cement is equivalent to 55 bricks (compared with 125 commercial bricks per bag of cement). However, the water consumption was calculated to be 16.5 litres per bag of cement, which is less than the 20 litres used in commercial brickmaking. Water is an expensive commodity in South Africa and using tailings to make the bricks saves water. Thus, the more economical option would be the second mixing ratio, since it uses less water and cement but still results in relatively high brick compressive strengths. Even though the second option is economically acceptable, the high cement content is a disadvantage.

However, regarding the overall brickmaking process some other factors should be considered. The brickmaking plant can be close to the tailings dumps in order to cut down on costs. In addition, it is important to note that most of the tailings material already occurs in fine form, therefore not much size reduction (which is an energy-intensive process) is required.

Since the use of tailings for brickmaking conserves natural resources, one could say that the benefit to the environment outweighs mere economic considerations. The use of tailings would mean that the companies have to spend less on waste management, while at the same time reducing human exposure to tailings, consequently reducing the effect that mine waste has on the health of inhabitants in the mining area. The use of gold mine tailings for brickmaking also constitutes an additional source of revenue for the gold mining companies and in the process creates jobs.

Conclusions and recommendations This laboratory-scale study was aimed at utilizing Witwatersrand gold mine tailings in making bricks. The results from XRD and XRF showed that the chemical composition of the Witwatersrand gold mine tailings is similar to that of the clay material used for commercial brickmaking. It was then concluded that it would be technically viable to use the tailings for brickmaking.

Following the South African masonry standards for brickmaking and testing, it was found that the commercial bricks have an average compressive strength of 750 kN and that the strongest bricks made from the tailings gave an average compressive strength of 530 kN. Results from water absorption tests showed that water absorption is higher in neutral solutions compared to acidic solutions. The rate of absorption is high in the first day, but then stabilizes. The weight loss over a seven-day period was negligible at 0.06%.

It is recommended that more tests be conducted with a wider range of tailings to cement ratios as this might lead to identifying a ratio that yields a stronger brick than what has been observed in this project. In addition, the sizes of the tailings used as aggregate should be varied to a wider range.

This can be achieved by adding overburden to the fine tailings material.

As regards the economic considerations, the tailings bricks were found to utilize more cement than the commercial

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bricks, possibly due to lack off plasticity in the tailing materials used. This is a disadvantage since cement is expensive. It is thus recommended that cheaper alternative additives that have a high plasticity or binding properties be explored in the place of cement. Looking at the bigger picture, the use of tailings as brickmaking material would have great advantages in terms of environmental conservation and reduction of waste management costs.

Since the XRD analysis showed that uranium is present in Witwatersrand gold tailings, extensive research with regard to the chemical properties and the chemical stability of the bricks produced from gold mill tailings is required.

Acknowledgments The authors would like to acknowledge the School of Chemical and Metallurgical Engineering, University of the Witwatersrand, Johannesburg for granting the first author the opportunity to complete her Bachelor’s degree in Metallurgical and Materials Engineering. In addition they would like to acknowledge all the laboratory personnel in the School who provided unlimited support during the research work.

Further, the authors would like to acknowledge the team of academics and laboratory personnel at the School of Civil Engineering for their guidance on masonry standards and for providing laboratory space and equipment during the project.

Lastly, much appreciation is due to AngloGold Ashanti for the idea behind the project and for supplying the tailings used in this study.

References AHMAD, S. and SAIFUL AMIN, A.F.M. 1998. Effect of curing conditions on the compressive strength of brick aggregate concrete. Journal of Civil Engineering, g vol. CE 28, no. 1. pp. 37–49.

AHMARI, S. and ZHANG, L. 2012. Production of eco-friendly bricks from copper mine tailings through geopolymerization. Construction and Building Materials, vol. 29. pp. 323–331.

ALGIN, H.M . and TURGUT, P. 2008. Cotton and limestone powder wastes as brick material. Construction and Building Materials, vol. 22, no. 6. pp. 1074–1080.

AMERICA'S CEMENT MANUFACTURERS. 2014. Curing. http://www.cement.org/ cement-concrete-basics/working-with-concrete/curing [Accessed 2 October 2014].

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BENNET, J.M., SUDHAKAR, M., and NATARAJAN, C. 2013. Development of coal ash GGBS based geopolymer bricks. European International Journal of Science and Technology, vol. 2, no. 5. pp. 133–139.

CHEN, Y., ZHANG, Y., CHEN, T., ZHAO, Y., and BAO, S. 2011. Preparation of ecofriendly construction bricks from hematite tailings. Construction and Building Materials, vol. 25. pp. 2107–2111.

CHOU, M.I., CHOU, S.F., PATEL, V., PICKERING, M.D., and STUCKI, J.W. 2006.

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CITY OF JOHANNESBURG (COJ). Mayoral Committee, 2013. 63 Amendment of tariff charges for water services and sewage and sanitation services:2013/14.

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DAILY TELEGRAPH. 2014. The brick directory words and blog. http://www.brickdirectory.co.uk/html/brickdirectorywords.html [Accessed 11 October 2014].

DEAN, K.C., FROISLAND, L.J., and SHIRTS, M.B. 1968. Utilisation and stabilisation of mineral wastes. Bulletin 688. US Bureau of Mines.

ELJHD ADMIN. 2011. An action plan for acid mine drainage. http://www.earthlife.org.za/2011/04/an-action-plan-for-acid-minedrainage [Accessed 10 October 2014].

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The viability of using the Witwatersrand gold mine tailings for brickmaking ENVIRONMENT. 2011. Acid mine drainage South Africa. f http://www.environment.co.za/poisoning-carcinogens-heavy-metalsmining/acid-mine-drainage-amd-south-africa.html [Accessed 10 October 2014].

FAST ONLINE. 1995. Global overview of construction technology trends: Energy efficiency in construction. http://www.fastonline.org/CD3WD_40/CD3W/ CONSTRUC/H1661E/EN/B856_6.HTM [Accessed 10 October 2014].

UNIVERSITY OF HIROSHIMA. 2001. Geology and non-fuel mineral deposits of Africa and the Middle East. http://www.home.hiroshima-u.ac.jp/er/Rmin_kS_ 01_A%26ME.html [Accessed 10 October 2014].

HYDRAFORM. 2014. Brick making. http://www.hydraform.com/PDFs/ vibraform/V2-LE.pdf [Accessed 19 October 2014].

JAIN, S.K., GARG, S.P., and RAI, M. 1983. Autoclaved calcium silicate bricks from gold mine tailings. Research and industry, vol. 28. pp. 170–172.

JONES, M.Q. 2003. Thermal properties of stratified rocks from Witwatersrand gold mining areas. Journal of the South African Institute of Mining and Metallurgy, vol. 103, no. 3. pp. 173–186.

KOUMAL, G. 1994. Method of environmental clean-up and producing building material using Cu mine tailing waste material. US patent US5286427 A. KREH, R.T. 2003. Masonry Skills. 5th edn. Delmar Learning, USA. LIU, Z., CHEN, Q., XIE, X., XUE, G., DU, F., and NING, Q. 2011. Utilization of the sludge derived from dyestuff-making wastewater coagulation for unfired bricks. Construction and Building Materials, vol. 25, no. 4. pp. 1699–1706.

MANOHARAN, C., SUTHARSAN, P., and DHANAPANDIAN, S. 2012. Characteristics of some clay materials from Tamilnadu, India, and their possible ceramic uses. Ceramica, vol. 58. pp. 412–418.

MINING FACTS. 2014. How are waste materials managed at mine sites. http://www.miningfacts.org/environment/How-are-waste-materialsmanaged-at-mine-sites [Accessed 25 September 2014].

MORCHHALE, R.K., RAMAKRISHNAN, N., and DINDORKAR, N. 2006. Utilization of copper mine tailings in production of bricks. Journal of the Institute of Engineers, Indian Civil Engineering Division, vol. 87. pp. 13–16.

MUELLER, H., MAITHY, S., and PRAJAPUTI, S. 2008. Greenbrick making manual.

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ROSNER, T. and VAN SCHALKWYK, A. 2000. The environmental impact of gold mine tailings foot prints in the Johannesburg region, South Africa. Bulletin of Engineering, Geology and Environmetal Studies, vol. 59. pp. 137–148.

ROY, S., ADHIKARI, G.R., and GUPTA, R.N. 2007. Use of gold mill tailings in making bricks: a feasibility study. Waste Management Research, vol. 25. pp. 475–482 SAEED, A. and ZHANG , L. 2012. Production of ecofriendly bricks from Cu mine tailings through geopolymerization. Construction and Building materials, vol. 29. pp. 323–331.

STATISTICS SA. 2013. South African statistics 2013. Johannesburg. 2013.

VERMEULEN, N.J. 2001. The Composition and State of Gold Tailings. University of Pretoria ETD. Retrieved 12 April 2014. Witwatersrand gold ore composition. http://upetd.up.ac.za/thesis/submitted/etd-03102006122937/unrestricted/02chapter2.pdf [Accessed 12 April 2014].

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ZHANG, L. 2013. Production of bricks from waste materials - a review.

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MATIVON BRICKS. 2012. Brick depot and building supplier. http://www.mativonbricks.co.za/products [Accessed 19 October 2014].

YANG, Y., ZHU, S., LI, Q., YANG, B., and CHEN, Y. 2011. Research on making fired bricks with gold tailings. International Conference on Computer Distributed Control and Intelligent Environmental Monitoring, g Qingdao Shandong, China.

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http://dx.doi.org/10.17159/2411-9717/2015/v115n4a9 ISSN:2411-9717/2015/v115/n4/a9

Evaluation of some optimum moisture and binder conditions for coal fines briquetting

by P. Venter* and N. Naude* Paper written on final year project work carried out in partial fulfilment of B.Eng degree in Coal Beneficiation

Coal mining is a thriving industry and 53% of the coal mined in South Africa is used for electricity generation. Mechanization has made coal mining more efficient, but fines generation has subsequently increased. Up to 6% of the run of mine material can report to the -200 μm fraction.

Common problems associated with fines handling include dust formation, storage problems, and high moisture levels. A method to turn this material into a saleable product instead of stockpiling it can add value to a company.

Briquetting is a pressure agglomeration method where loose material is compacted into a dense mass (FEECO International, 2014). The briquettes must be able to withstand rigorous handling and transport operations without disintegrating. This study aims to investigate the optimum binder and moisture conditions required to produce a mechanically strong briquette using two different binders – a PVA powder (binder A) and a starch powder (binder B).

It was found that for binder A the optimum moisture level was 12% to 14%. At this moisture level the greatest compression strength gains were observed, and low amounts of fines produced in impact and abrasion tests.

The minimum amount of binder added while still obtaining a strong briquette was 0.5% binder A. For binder B the optimum moisture level was also 12% and the minimum amount of Binder B to be added was found to be 1%. Briquettes that were dried outside reached their peak strength after about four days, whereas the briquettes that dried inside took about 20 days to reach their strength plateau. Hardly any degradation took place on the surface of the binder A film after exposure of 300 hours of artificial weathering. Thermogravimetric analysis confirmed that neither binder A nor binder B will add to the ash content of the coal fines, as both binders totally decompose above 530°C. Binder B yielded stronger briquettes after 15 days and also generated less fines. It is therefore superior to binder A and would be recommended for further use.

Keywords coal fines, briquetting, binder, moisture level.

Introduction The need for coal briquetting The coal mining industry in South Africa has been operating for more than 120 years.

Annually, 224 Mt of coal is produced, of which 25% is exported. The remainder is used to feed South Africa's industry, with 53% used for generating electricity (Eskom, 2014). Electricity generated from coal-fired power stations accounts for 77% of South Africa's electricity supply.

Fines generation in coal mining has increased as a result of mechanized mining, and up to 6% of the run of mine product can be in the -200 µm fraction. Coal-fired power stations The Journal of The Southern African Institute of Mining and Metallurgy

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Synopsis

in South Africa do not accept a product of -200 µm size because of the high moisture content(England, 2013). In this paper, any -8 mm material will be considered as fines.

Additional problems arising from coal fines generation include flow problems from containers, dust formation in plants and fire hazards during stockpiling.

An increased moisture content inevitably reduces the calorific value of the coal, as well as increasing handling problems. Instead of pumping these fines to slime dams or discarding them in old workings, a means of economical agglomeration can be beneficial to mines and power stations.

ESI Africa (2014) estimates the amount of thermal-grade fines stockpiled over the past 100 years at about 1 billion tons. Only in the last few years have methods to utilize these stockpiles in South Africa started to be explored. These methods include briquetting, pelletizing, and granulation.

Briquetting is a pressure agglomeration method where loose material is compacted into a dense mass (FEECO International, 2014). It is a more advanced and more expensive process than pelletization and granulation, but the endproduct can withstand the rigorous handling methods that export coal undergoes, and in some cases, is water-resistant.

Table I shows that, assuming that Eskom will pay a similar price per ton of briquettes to that of coarse coal, it is not financially viable to produce briquettes using a generous amount of binder. Producing briquettes with a modest amount of binder look more promising.

According to Sastry et al., binder cost may represent 60% of the total cost of briquette manufacture. However, since low binder additions can be detrimental to the mechanical strength of the briquettes, a compromise must be reached between binder content and mechanical strength.

Evaluation of some optimum moisture and binder conditions for coal fines briquetting Table I

Binder costs per ton of briquettes Cost (R/ton) Coal (selling price) Minimum Maximum

Binder A Binder A (0.1 wt%) (0.9 wt%)

150 400

Binder B Binder B (0.3 wt%) (3 wt%)

The fines should have a low moisture content, good compatibility, small particle size, and a wide particle size distribution to facilitate good packing of the particles. Proper mixing is also critical to ensure that the binder is distributed evenly throughout the mixture.

Physical testing of briquettes 20 28

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1 3

149 298

A review of previous work done on briquetting The earliest coal briquettes were made in hand-filled brick moulds using clay and cow dung as binders. These bricks had poor mechanical properties which made them unsuitable for transportation over long distances. Only by the 1850s were mechanical methods introduced to briquette brown coal and lignites without the use of binders, while hard coal briquettes required binders to stay intact.

Roller presses were first developed in Belgium by Louiseau to address the need for strong briquettes. Pillow-shaped holes in the roller faces compact material into dense briquettes that weighs no more than 50 g each. The basic principles of these machines have remained relatively unchanged over the years apart from small improvements to extend the life and reduce maintenance.

The briquetting process is conducted at room temperature and the use of binders depends on the coal grade being used.

The complete briquetting process and binder options are discussed later in this paper.

Binders for coal briquetting It is possible to use binderless pelletization for coal; however, most coal fines are not self-agglomerating. The literature suggests numerous binders for the pelletization of coal (Altun et al., 2001; Dehont, 2006): ® Coking and oil refining residues such as tar and coal pitch ® Residues from paper mills (lignosulphonate) ® Molasses with possible additions of lime ® Starch ® Synthetic resins ® Synthetic polymers .

These binders should have adequate binding strength, relatively low cost, and be resistant to weathering (Waters, 1969).

According to Richards (1990), the most important physical properties of briquettes are resistance to crushing, impact, abrasion, and water penetration. These properties are all heavily dependent on the development of strong and durable bonds between particles during the agglomeration stage. It is critical that briquettes withstand storage, handling, and transport during which they will be dropped, experience abrasion on conveyors, and be exposed to the elements. Four laboratory tests are recommended to monitor the physical strength properties of briquetted fuels either during process development or commercial production. These are a drop shatter test, a crushing resistance test, a tumbler abrasion test, and an immersion water resistance test (Richards, 1990).

Method Coal fines from Exxaro's Mafube coal mine near Middelburg were used in the briquetting process. This study forms part of research by Exxaro into coal fines agglomeration, and the binders that were used were prescribed by Exxaro. Two binders were used – a PVA powder (binder A) and starch powder (binder B). Coal fines from Exxaro's Mafube coal mine near Middelburg were used in the briquetting process.

Ten days of testing were allocated for physical tests, ranging from day 0 to say 35. ‘Day 0’ refers to the day that the briquettes were manufactured, as these briquettes have not been allowed to dry for a full day.

The mechanical properties of the briquettes were investigated by means of following tests.

Compressive strength Compressive strength is the maximum crushing load a briquette can withstand before cracking or breaking. A single briquette was placed on the platform of the tensile strength testing machine and, with the machine operating in the compressive mode, a constant load was applied until the briquette fractured.

The load at fracture can also be converted to a stress using the equation:

Briquetting process The main operations during a briquetting process are as follows (Waters, 1969): ® Screening and drying of the coal (if too wet) ® Mixing of coal with binder ® Feeding to briquette machine and pressing ® Drying ® Storing and packaging.

According to Dehont (2006), the size distribution of the coal fines should be roughly as follows: ® 50% from 0 to 0.5 mm ® 25% from 0.55 to 1 mm ® 20% from 1 to 2 mm ® 5% from 2 to 3 mm.

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By expressing the load as a stress (force per unit area) it is possible to compare briquettes of various sizes and incorporating different binders. For the purpose of this study all briquettes were of similar size and shape, and therefore only the load force was used. A batch of 20 briquettes was tested at a time.

Impact resistance A batch of 20 briquettes was dropped once from a height of 2 m, and the particle size analysis of the pellets and broken pieces conducted. According to Richards (1990), ‘Impact resistance testing is considered to be the best general diagnostic of briquette strength’.

The Journal of The Southern African Institute of Mining and Metallurgy

Evaluation of some optimum moisture and binder conditions for coal fines briquetting Abrasion resistance A charge consisting of 20 briquettes was rotated in a tumbler machine for 100 revolutions at 50 revolutions per minute. The tumbler drum had dimensions of 278 mm in length by 20 mm in diameter, and a 38 mm wide lifter plate was also welded along the length of the drum. The charge was then collected and the particle size distribution (PSD) of the fines (-8 mm) that were generated was calculated.

Water resistance Since briquettes may in some cases be stockpiled outside and exposed to the elements, it is important to test for water resistance. A single weighed briquette was submerged in a beaker of cold tap water and inspected for disintegration by applying finger pressure at intervals of about ten minutes. If the briquette remained intact after 30 minutes, the surface water was wiped off with a cloth and the briquette was weighed again. To obtain a quantitative comparison, a water resistance index (WRI) was calculated as follows: WRI = 100–%water after 30min Richards (1990) argues that a WRI > 95% should be obtainable after 30 minutes

Artificial weathering of binder A A QUV Accelerated Weathering Tester was fitted with A340 UV lamps. Binder A was exposed to alternating cycles of UV light and elevated temperatures. The temperature was set to 63°C and the irradiance at 0.67 W/m2, and the samples were run on a dry cycle.

After exposure of only a few days the QUV tester can reproduce damage that will take months or years outdoors.

The rate of polymer oxidation was measured by conducting infrared spectroscopy (IR) on the exposed films. By following the growth of the carbonyl peak near 1720 cm-1, a carbonyl index was defined by the ratio of this absorption to that at 2900 cm-1 and used to quantify the progression of degradation.

fines. Briquettes with 12% moisture showed the greatest strength gains for both binder A and binder B. Day 0 compression strength results could not be obtained from briquettes with 18% moisture and using binder B, as they were too soft and disintegrated under the load.

In Figure 2 the strong dependence of compressive strength on binder content is illustrated very clearly. The initial strength, as well as day 15 strength, increases with increase in binder content. Briquettes with binder B at levels of 0.3% and 0.5% show poor strength, and attain strength values well above the SABS standard of 25 N only with binder additions of 1% and more. The maximum strength of 358 N was obtained on day 15 for 3% binder B. However, it may be necessary to make a trade-off between the strength and binder cost. Lower additions of binder B result in acceptable briquette strengths, with gains of 61 N and 97 N for 1% and 2% binder B addition respectively.

Low levels of binder A also resulted in weak briquettes.

Only at a level of 0.5% binder A and higher is sufficient compression strength achieved in the briquettes. The highest binder A addition of 0.9% also resulted in the greatest strength gain of 68 N. This is followed by 0.5% binder A with a gain of 46 N, which will be a more economical option.

The observed trend of increasing compressive strength with increased binder addition is expected because at higher binder levels more binder is available to be dispersed between the coal particles and ensure bonding between binder and coal.

Thermogravimetric analysis The sample was heated to a maximum temperature of 900°C, and the residual mass plotted against temperature. The material that remains after the maximum temperature is reached should correspond to the total ash percentage of the coal sample.

Results and discussion

Figure 1 – Compressive strength as a function of varying moisture content (constant binder addition)

Compressive strength

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Figure 2 – Compressive strength as a function of binder addition (constant 12% moisture) VOLUME 115

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The SABS 1399:1999 standard was used to ensure the briquettes met the minimum compressive strength requirements. This standard specifies the requirements for charcoal made from wood in either lump or briquette form. The compression strength of the briquettes was first evaluated as a function of moisture content (the total moisture of the batch during the briquetting process). From Figure 1 it can be seen that the compressive strength increased from day 0 to day 15 for each moisture level. At 10% moisture level, briquettes with binder B, with an initial strength of 37 N, gained hardly any strength after 15 days. Binder A yielded an increase of only about 40 N after 15 days. This low moisture content does not allow for efficient dispersion of the binder throughout the coal

Evaluation of some optimum moisture and binder conditions for coal fines briquetting Drop and tumble tests The results of the drop and tumble tests were combined on a single axis for easier comparison. In Figure 3 the binder addition is constant, with moisture as the variable. SABS standard 1399 specifies a maximum of 5% fines. Fines are considered to be material that passes through an 8 mm screen.

The tumble test produced more fines than the drop test for all moisture contents. For day 0 testing, when the briquettes are at their weakest, the briquettes with binder A generated on average 24% fines, while for binder B this was significantly less at 8%.

For both binders, the greatest amount of fines generated during both the drop test and the tumble test were from the briquettes with lowest moisture level of 10%. This is probably a result of the inefficient distribution of the binder throughout the mixture.

The amount of fines generated in both tests decreased with increasing moisture level, but this trend was more pronounced with the drop test. For binder A the optimum moisture level indicated by drop test and tumble test results is 16% and 18% respectively. For binder B, a moisture level anywhere between 12% and 18% will give similar results from drop tests and tumble tests.

In tests where the binder content was varied and the moisture level was kept constant at 12%, the drop test and tumble test results followed similar trends as in Figure 3, with fewer fines being produced in the drop test. The optimum

amount off binder A was found f to be 0.5%. This amount produced the second lowest amounts of fines during the tumble test and drop test, at 8.8% and 32.6% respectively.

From an economic point of view, adding 0.5% binder is preferable to adding 0.9%.

An amount of 1% and higher of binder B will give favourable results. Very small amounts of fines were generated during the tumble tests for 1%, 2%, and 3% binder additions.

This is an indication of how well binder B briquettes can withstand an abrasive environment like transport on a conveyor belt. Hardly any fines were generated for the three higher amounts of binder additions, and hence these briquettes will stay relatively intact when tipped from a transport truck or when falling from a conveyor belt.

The optimum binder B content is therefore 1%, being the level that will ensure a low amount of fines generated.

Drying conditions Figure 5 shows the strength gain for both binders over the first five days of drying under different conditions. Drying outside in direct sunlight exposes the briquettes to a higher average temperature than those allowed to dry indoors. This results in faster evaporation of the moisture and the different bonds created by the binders are established much earlier in the curing process. For both binders, a strength plateau is reached after five days of drying outside, and little additional strength is gained from day 5 to day 35.

For the samples dried indoors, the greatest strength gain is achieved in the first two days of curing. At this stage these briquettes do not have the same strength as the samples dried outside, but from day 5 they continue to steadily gain strength until the final strength on day 35 is similar, or close to, that of the samples that were dried outside.

Drying samples outside looks like the obvious choice as this allows for a much shorter curing process. However, it should be noted that neither of the binders results in a waterresistant briquette, and therefore when these briquettes are stored outside, they must be under cover to prevent rain damage.

Water resistance

Figure 3 â&#x20AC;&#x201C; Drop test and tumble test results for briquettes with varying moisture (constant 0.5% binder A)

Figure 4 â&#x20AC;&#x201C; Compressive strength of briquettes with binder A and binder B and 12% moisture that were dried inside and outside

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When binder B briquettes were submerged in water, they immediately disintegrated. It was clear that these briquettes did not have any water resistance.

Figure 5 â&#x20AC;&#x201C; Infrared spectra of new and weathered binder A film

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Evaluation of some optimum moisture and binder conditions for coal fines briquetting

Figure 6 – Change in mass of coal and binders during thermogravimetric analysis

Binder A was expected to impart water-resistant properties to the briquettes, but these also disintegrated when submerged in water. Even after 15 days of curing the binder was not able to establish water-resistant bonds between the particles.

For binder B the optimum moisture level was also 12%, and the minimum binder addition was found to be 1%.

The briquettes that were dried outside reached their peak compressive strength after about four days. The briquettes that dried inside took about 20 days to reach maximum strength.

Neither of the binders resulted in water-resistant briquettes, as all of the briquettes tested disintegrated when submerged in water.

ATR spectroscopy indicated that no degradation of the binder A film took place after 300 hours of exposure in the QUV. The TGA results confirmed that neither binder A nor binder B will increase the ash content of the coal fines, as both binders totally decompose above 530°C. The cost of binder B is higher than binder A, but its strength after 15 days of curing and the low amount of fines produced with minimum of 1% binder addition makes it the preferred binder to use.

References

Artificial weathering of binder A The results from attenuated total reflection (ATR) spectroscopy indicated that no weathering took place on the surface of the finder A film. In Figure 5 barely any difference is seen between the curves for the different exposure times, indicating that little to no degradation had taken place on the surface of the film after exposure of 300 hours in the QUV.

ALTUN, N. E., HICYILMAZ, C., and KÖK, M.V. 2001. Effect of different binders on the combustion properties of lignite. Part 1. Effect of thermal properties. Journal of Thermal Analysis and Calorimetry, vol. 65. pp. 787–795.

DEHONT, F. 2006. Coal briquetting technology. http://www.almoit.com/allegati/applicazioni_particolari/15/COAL%20BRIQU ETTING%20TECHNOLOGY.pdf [Accessed 13 October 2013].

Thermogravimetric analysis

Conclusions The compressive strength of the briquettes depends on the binder addition and the moisture content. For both binder A and binder B, the optimum moisture level was 12%. The minimum binder addition for adequate strength was 0.5% for binder A, and 1% for binder B. Further additions of binder B increased briquette strength, but the higher cost of binder B renders this option uneconomical.

For binder A, the optimum moisture level was 12% to 14%.

At this moisture level the largest compressive strength gains were observed, as well as a low amount of fines produced. The minimum amount of binder to be added to obtain a briquette of adequate strength was 0.5%.

The Journal of The Southern African Institute of Mining and Metallurgy

ENGLAND, T. 2013. The economic agglomeration of fine coal for industrial and commercial use: A review of past and present work both locally and internationally. http://www.coaltech.co.za/chamber%20databases%5Ccoaltech%5CCom_Doc Man.nsf/0/09EFEE68D9257C164225781B00364F21/$File/Task%204.4%2 01%20-%20Agglomoration%20of%20Fine%20Coal%20%20Trevor%20England.pdf ESI-AFRICA. 2014. The economic possibilities of South Africa’s coal fines. http://www.esi-africa.com/the-economic-possibilities-of-south-africas-coalfines/ [Accessed 17 June 2014].

ESKOM. 2014. Coal Power. http://www.eskom.co.za/AboutElectricity/ElectricityTechnologies/Pages/Coa l_Power.aspx [Accessed 17 June 2014].

FEECO INTERNATIONAL. 2014. Briquettes, Granules, and Pellets – What’s the difference? http://feeco.com/2012/01/11/briquettes-granules-and-pelletswhats-the-difference/ [Accessed 17 June 2014].

RICHARDS, S.R. 1990. Physical testing of fuel briquettes. Fuel Processing Technology, vol. 25. pp. 89–100.

SABS 1399:1999. Wood charcoal and charcoal for household use. South African Bureau of Standards. Pretoria. http://www.sabs.co.za [Accessed 20 November 2014].

SASTRY, K.V.S. and FEURSTENAU, D.W. 1977. Kinetics and process analysis of agglomeration of particulate materials. Agglomeration '77.

7 AIME, New York. pp. 318-402.

WATERS, P.L. 1969. Binders for fuel briquettes: a critical survey. Technical Communication 51. CSIRO. VOLUME 115

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The results from TGA are plotted in Figure 6. Hardly any difference is seen between the pure coal and coal with binder mixtures. The coal, with or without binder, is thermally stable up to 400°C, and between 400°C and 600°C the binder and other volatiles decompose. Above 600°C only about 30% of the material remains, and this value is similar to the total ash content of 30.4% shown in Figure 6.

Binder A starts decomposing at about 95°C and is totally decomposed at about 530°C. Binder B shows slow decomposition from 25°C to 90°C followed by rapid decomposition between 90°C and 490°C. Binder B is totally burned off above 500°C. The binder additions to the coal were the maximum amounts that were used throughout this study: 0.9 % binder A and 3% binder B. It is safe to say that neither of the two binders will add to the ash content of the coal as both binders totally decompose above 530°C.

Copper Cobalt Africa

In association with The 8th Southern African Base Metals Conference

6–8 July 2015 Zambezi Sun Hotel, Victoria Falls Livingstone, Zambia

For further information contact: Head of Conferencing Raymond van der Berg, SAIMM P O Box 61127, Marshalltown 2107 Tel: +27 (0) 11 834-1273/7 E-mail: raymond@saimm.co.za Website: http://www.saimm.co.za

Second Announcement & List of Abstracts

Join us for the inaugural Copper Cobalt Africa Conference in the heart of Africa.

To be held at Victoria Falls, one of the Seven Natural Wonders of the World, this prestigious event will provide a unique forum for discussion, sharing of experience and knowledge, and networking for all those interested in the processing of copper and cobalt in an African context, in one of the worldʼs most spectacular settings.

The African Copper Belt has experienced a huge resurgence of activity in recent years following many years of political and economic instability. Today, a significant proportion of capital spending, project development, operational expansions, and metal value production in the Southern African mining industry are occurring in this region. The geology and mineralogy of the ores are significantly different from those in other major copper-producing regions of the world, often having very high grades as well as the presence of cobalt. Both mining and metallurgy present some unique challenges, not only in the technical arena, but also with respect to logistics and supply chain, human capital, community engagement, and legislative issues. This conference provides a platform for discussion of these topics, spanning the value chain from exploration, projects, through mining and processing.

For international participants, this conference offers an ideal opportunity to gain in-depth knowledge of and exposure to the Southern African base metals industry, and to better understand the various facets of mining SPONSORS: and processing in this part of the world that both excite and frustrate the industry.

Premium A limited number of places are available for post-conference tours to Zambiaʼs most important commercial operations, including Kansanshi, the largest mine in Zambia, with 340 kt/y copper production and its soon-to-be-completed 300 kt/y smelter, and Chambishi Metals.

Jointly hosted by the mining and metallurgy technical committees of the Southern African Institute of Mining and Metallurgy (SAIMM), this conference aims to: • Promote dialogue between the mining and metallurgical disciplines on common challenges facing the industry, • Encourage participation and build capacity amongst young and emerging professionals from the Copper Belt region, • Improve understanding of new and existing technologies, leading to safe and optimal resource utilisation.

The organising committee looks forward to your participation.

http://dx.doi.org/10.17159/2411-9717/2015/v115n4a10 ISSN:2411-9717/2015/v115/n4/a10

Air drying of fine coal in a fluidized bed

by M. Le Roux*, Q.P. Campbell*, M.J. van Rensburg*, E.S. Peters*, and C. Stiglingh* Paper written on project work carried out in partial fulfilment of Degree in Chemical Engineering (NWU) — Pursuing Masters in Coal Beneficiation

The demand for energy has continued to rise worldwide in line with population growth. The majority of South Africa’s electricity is supplied by coal-fired power stations. The amount of fine coal (-2 mm) generated at coal processing plants has increased, due mainly to mechanized mining methods. Fine coal retains more water, which lowers its heating value.

Drying the coal is costly and it is difficult to achieve the required moisture content. Consequently, coal fines are often discarded. An estimated 8% of the total energy value of mined coal is lost1.

Fluidized bed technology is often used to dry coal thermally, but this method is expensive and has an adverse environmental impact. The objective of this study was to investigate the removal of moisture from fine coal (<2 mm) in a fluidized bed operated with dry fluidizing air at moderate temperatures as the drying agent. The effects of different air temperatures and relative humidity levels were investigated in a controlled environment.

The study further investigated the influence of coal particle size on moisture removal.

The drying rate was found to increase with increasing temperature. The relative humidity of the drying air had a more pronounced effect on the drying rate, even at temperatures as low as 25°C.. It became more challenging to remove moisture as the particle size decreased. The gain in calorific value was greater than the energy required to dry the coal samples, showing that a fluidized bed using moderately warm dry air is an energy-efficient drying technology. The energy efficiency of the fluidized bed compared favourably with other thermal drying methods.

Keywords coal fines, drying, fluidized bed, energy efficiency.

Introduction The increased use of mechanized coal mining methods has resulted in greater amounts of coal fines being generated. Many operations report an estimated 6% of their ROM production to be in the -2 mm size range. Fine and ultra-fine size ranges constitute about 11% of the nominal product and retain the bulk of the moisture (SANEDI, 2011). Given the problems associated with moisture in fine coal, it is important to investigate and improve available moisture removal techniques. Coal constitutes the primary source of energy in South Africa and is a major contributor to the economy, and therefore improving coal quality will in effect

1

Rowan, S.L. 2010. Analysis and scaling of a two-stage fluidized bed for drying of fine coal particles using shannon entropy, thermodynamic exergy and statistical methods. PhD dissertation, University of West Virginia, Morgantown WV.

The Journal of The Southern African Institute of Mining and Metallurgy

maximize the quantity of usable coal (De Korte and Mangena, 2004). With the use of effective dewatering methods, fine coal can be beneficiated and added to the coarse particle circuits without compromising the quality of the product and hence substantially increasing the overall plant yield (Condie and Veal, 1998). Condie and Veal (1998) suggest that by rule of thumb, for every ton of moisture removed the clean coal product stream is supplemented with about 4 t of fine coal. This is a powerful incentive for developing advanced dewatering techniques for fine coal particles. Additionally, excessive moisture adds to the mass-based transport costs of coal. For this reason, developing advanced efficient fine coal drying techniques is beneficial from an economical point of view (Campbell, 2006).

Rowan (2010) states that coal preparation plants generally discard coal fines with size fractions below 150 μm into waste ponds. This poses a danger of spontaneous combustion, acid mine drainage, and dust release as the surface of the coal is exposed to ambient air and weathering conditions for long periods (De Korte and Mangena, 2004). Coal fines are more susceptible to water absorption than coarser coal and can contain up to 25 wt% total moisture after filtration (Le Roux, 2003).

Thermal drying methods are more efficient than mechanical dewatering techniques (De Korte and Mangena, 2004), but the price of coal limits the use of these methods (SANEDI, 2011).

Studies conducted at North-West University showed that drying of fine coal (-2 mm +1 mm) in a fluidized bed is possible at low temperatures between 25°C and 40°C. Work by Le Roux et al. (2012) on vacuum filtration showed that intentionally damaging a filter cake improved the airflow infiltration, leading to a lower pressure differential across the cake but increasing the dewatering

* School of Chemical and Minerals Engineering, North-West University, Potchefstroom, South Africa. © The Southern African Institute of Mining and Metallurgy, 2015. ISSN 2225-6253. Paper received Jan. 2015 VOLUME 115

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Synopsis

Air drying of fine coal in a fluidized bed the efficiency ff off drying fine f coal particles in a fluidized f bed.

The energy consumption during the drying process was calculated and compared to published data on thermal drying processes.

Sample preparation

Figure 1 – Forms of moisture related to coal (after Lemley et al., 1995)

efficiency. This work confirmed that high airflow conditions resulted in a lower final cake moisture content of 3–5 wt%.

Further studies found that an increased airflow rate resulted in a more effective moisture transfer from the coal fines to the drying air (Le Roux et al., l 2013).

The moisture content of fine coal particles is made up of surface, capillary, and chemically bound moisture (Rong, 1993) as depicted schematically in Figure 1. Free moisture is found on the exterior surface of coal particles (Condie and Veal, 1998), and can be removed by mechanical methods such as filters and centrifugal units. Capillary-bound moisture is absorbed and held tightly within micro-capillaries and micropores of individual coal particles (Rong, 1993). Removal of this moisture calls for thermal drying techniques for complete drainage (Condie and Veal, 1998). Chemically bound moisture is not included when measuring the total moisture content of the coal (Campbell, 2006), and can be removed only by pyrolysis.

The equilibrium moisture content of coal is characterized as the moisture content at which the coal particles no longer gain or lose moisture, and it varies according to the temperature and relative humidity conditions of the atmosphere surrounding the particles. Mechanical methods are insufficient for the removal of this equilibrium moisture, which can be reduced only by means of evaporation (Le Roux et al., l 2013). The relative humidity and temperature act as driving forces that change the phase equilibrium between vapour and liquid, with lower humidities and higher temperatures leading to moisture being absorbed from the particle by the drying medium (Koretsky, 2004).

South African bituminous coal from the Waterberg coalfield was used for these experiments. The proximate analysis of this type of coal is given in Table I. The coal was crushed and sieved into three particle size ranges: fines (between 2 mm and 1.18 mm) and ultra-fines (between 1.18 mm and 0.5 mm).

The samples were drenched in water for a day and the excess free moisture was removed by pressure filtration. The moisture content of each filtered sample was determined (SANS5925:2007) before the coal was fed to the fluidized bed for dewatering.

Apparatus A fluidized bed column (10 cm inner diameter × 40 cm length) was constructed from polycarbonate (Figure 2). The column was connected to a blower, which was used to draw conditioned air at a set temperature and relative humidity from a climate chamber (CTS climate test chamber Type: C-40/100).

A packed bed of glass marbles in the bottom section of the fluidized bed acted as airflow distributor. Mesh covers (0.5 mm aperture) were placed at the top and bottom sections of the fluidized bed to retain the bulk coal sample within the cylinder.

The outlet air from the column was returned to the climate chamber, and was recirculated to the column after the temperature and relative humidity values attained the pre-set levels. For each test, 100 g of fine coal sample with a total moisture content of approximately 25-35 wt% (typical of a pressure filter product) was fed to the fluidized bed cylinder.

The weight of the column was continually monitored during fluidization to determine the loss of moisture. A number of

Experimental method The aim of this project was to determine the effect of temperature, relative humidity, and particle size distribution on

Table I

Proximate analysis (air-dried basis)

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Component

Percentage by weight

Fixed carbon Moisture content Ash content Volatile matter

39.98 2.58 22.82 34.62

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Figure 2 – Experimental set-up of the fluidized bed (Van Rensburg, 2014) The Journal of The Southern African Institute of Mining and Metallurgy

Air drying of fine coal in a fluidized bed selected experiments were repeated in the fluidized f bed to determine the repeatability of the results.

Results and discussion Influence of temperature To study the effect of air temperature, wet samples of 100 g containing about 30 wt% total moisture were placed in the fluidized bed cylinder, and drying air was introduced at a superficial velocity of 1.5 to 1.7 m/s, which was slightly above the predetermined minimum fluidization velocity. Two sets of experiments were conducted at air temperatures of 25°C and 55°C respectively and a relative humidity (RH) of 30%.

Figure 3 shows the moisture loss from the ultra-fine sample (-1.18 mm +0.71 mm) under these conditions. The drying rate was quicker at 55°C than at 25°C. The drying time was reduced from 28 minutes to 21 minutes. This confirms the observation made by Rowan (2010) that elevated temperatures lead to higher dewatering rates. Higher temperatures disrupt the phase equilibrium and increase the amount of water transported from the coal sample into the surrounding air (Condie and Veal, 1998).

Duplicate experimental runs proved the repeatability of the results, the maximum and minimum standard deviation being 3.40 wt% and 0.12 wt% respectively.

Influence f off relative humidity For the next set of experiments, the fluidizing air was introduced at relative humidities of 30%, 50%, and 70% at a constant temperature of 55°C. The drying curves are shown in Figure 4. A comparison of Figure 4 with Figure 3 shows that relative humidity has a greater effect on the drying rate than temperature. Lower relative humidities counteract the capillary forces retaining the moisture in the coal particle, leaving a dried product in about 14 minutes at 30% RH, compared to over 30 minutes for 70% RH at the same temperature. Higher relative humidities weaken the moisture transfer mechanism, and therefore the moisture is displaced from the capillary channels at a lower rate (Condie and Veal, 1998). Van Rensburg (2014) stated that the mechanism for the transfer of water molecules from the coal to the air is enhanced when the drying air contains low moisture levels. This leads to a high transfer rate of the water molecules from an area of high moisture content to an area of low moisture content.

Influence of particle size Three wet filter cake samples, all with an initial moisture content of 33 wt%t and different size ranges (-1.18 mm +0.71 mm, -0.71 mm +0.50 mm, and a 50/50 mixture of -1.18 mm +0.71 mm and -0.71 mm +0.5 mm), were dried in the fluidized bed using feed air at 55°C and 50% RH. Figure 5 shows the drying curves for these experiments.

The coarse fraction (-1.18 mm +0.71 mm) and the 50% mixture showed similar drying responses, reaching a final moisture value in less than 20 minutes, while the finer fraction (-0.71 mm +0.5 mm) reached a similar final moisture value only after 37 minutes. This shows that it is increasingly difficult to remove moisture as the size of the coal particles decreases. Small particles have large surface areas and more micropores to absorb water, resulting in a higher degree of water retention (De Korte and Mangena, 2004).

Drying rates

Figure 4 – Effect of relative humidity on the drying of the mixture -1.18 mm +0.71 mm and +0.71 mm -0.5 mm at 55°C The Journal of The Southern African Institute of Mining and Metallurgy

Figure 5 Effect of particle size on the drying of coal at 55°C and 50% RH

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Figure 3 – Effect of temperature on the drying of -1.18 mm +0.71 mm coal at 30% RH

Figure 6 shows the different drying rates for fluidizing feed air conditioned at 25°C across the set relative humidity ranges. A decrease in relative humidity clearly increases the drying rate for all particle size ranges. The drying rate increased from 0.010 wt%/min to 0.015 wt%/min for the -2 mm +1.18 mm particles at 25°C with 70% and 30% RH respectively. It is also apparent that the drying rate increased with an increase in

Air drying of fine coal in a fluidized bed particle size at less than 50% RH. It is noteworthy that at 25°C, particle size is not the rate-limiting factor at relative humidity conditions exceeding 50% RH. The -0.71 mm +0.5 mm fraction had the lowest drying rate at 25°C and 70% RH, while the fastest drying rate was for the -2 mm +1.18 mm particle size range at 30% RH. It is clear that relative humidity has a more significant effect on the drying rate of the coal particles than temperature, hence dry air at moderately low temperatures is effective in drying fine coal.

Energy consumption The energy required during the drying process was calculated by considering the change in enthalpy of the water, the work done by the blower, as well as the energy required for the conditioning of the air. A temperature of 25°C and relative humidity of 50% were chosen as a basis for the calculations, since these were the average ambient conditions in the laboratory. The calorific value of the coal was upgraded by 8 MJ/kg on average using air at these conditions as drying medium. Figure 7 shows the calculated maximum and minimum of energy requirements to dry to three different moisture levels using the fluidized bed with those of other existing thermal drying technologies. It can be seen that fluidization is more energy-efficient than other thermal processes, with the exception of the Fleissner process.

Conclusion This work has shown that fine and ultra-fine coal particles can be dried at moderate temperatures and low relative humidities in a fluidized bed. The time required for fines is about half of that required to dry ultra-fines. The main driving force for removal of moisture from fine and ultra-fine coal is relative humidity. Energy calculations demonstrate that fluidization is more energy-efficient than other thermal drying processes.

Using dry air at a moderate temperature in a fluidized bed to dry coal particles is thus a promising technique warranting further study and development, since it has a potential energy advantage as well the ability to increase the calorific value of the coal.

Acknowledgements The authors would like to acknowledge the following institutions for their contribution towards this project: ® Coaltech ® NRF (National Research Foundation).

This work is based on research supported by the South African Research Chairs’ Initiative of the Department of Science and Technology and the National Research Foundation of South Africa. Any opinion, finding, or conclusion, or recommendation expressed in this material is that of the authors and the NRF does not accept any liability in this regard.

References CAMPBELL, Q.P. 2006. Dewatering of fine coal with flowing air using low pressure drop systems. PhD dissertation, North-West University, Potchefstroom. 130 pp.

CONDIE, D. and VEAL, C. 1998. Improved fine coal dewatering via modelling of cake desaturation. CSIRO, Australia. pp. 1–34.

DE KORTE, G.J. and MANGENA, S.J. 2004. Thermal Drying of Fine and Ultra-fine Coal. Report no. 2004 – 0255. Division of Mining Technology, CSIR, Pretoria. pp. 5–24.

KARTHIKEYAN, M., ZHONGHUA, W., and MUJUMDAR, A.S. 2009. Low-rank coal drying technologies – current status and new developments. Drying Technology, vol. 27, no. 3. pp. 403–415.

KORETSKY, M.D. 2004. Engineering and Chemical Thermodynamics. 2nd edn. John Wiley & Sons Hoboken, NJ. LE ROUX, M. 2003. An investigation into an improved method of dewatering fine coal. Master’s dissertation, North-West University, Potchefstroom. 96 pp.

Figure 6 – Drying rates of different size fractions at 25°C

LE ROUX, M., Campbell, Q.P., and Smit, W. 2012. Large-scale design and testing of an improved fine coal dewatering system. Journal of the Southern African Institute of Mining and Metallurgy, vol. 112, no. 7. pp. 673–676.

LE ROUX, M., CAMPBELL, Q.P., and VAN RENSBURG, M.J. 2013. Fine coal dewatering using high airflow. International Journal of Coal Preparation and Utilization, vol. 34. pp. 220–227.

SANEDI (South African National Energy Development Institute). 2011. http://www.sanedi.org.za/coal-roadmap/ [Accessed 25 June 2014].

RONG, R.X. 1993. Literature review on fine coal and tailings dewatering.

Advances in Coal Preparation Technology, vol. 2. Project P239A. JKMRC, University of Queensland. Brisbane. Australia. 120 pp.

ROWAN, S.L. 2010. Analysis and scaling of a two-stage fluidized bed for drying of fine coal particles using shannon entropy, thermodynamic exergy and statistical methods. PhD dissertation, University of West Virginia, Morgantown WV. 154 pp.

VAN RENSBURG, M.J. 2014. Drying of fine coal using warm air in a dense medium fluidised bed. Master’s dissertation, North-West University, Potchefstroom.

98 pp.

Figure 7 – Energy consumption of different thermal drying methods for coal in the -2 mm +1.18 mm particle size range

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YU, A.B., STANDISH, N., and LU, L. 1994. Coal agglomeration and its effect on bulk density. Powder Technology Journal, l vol. 82, no. 1. pp. 177–189.

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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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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

Website: http://www.saimm.co.za

N CONFERENCE Mining, Environment and Society Conference 12–13 May 2015, Mintek, Randburg, Johannesburg N CONFERENCE Copper Cobalt Africa Incorporating The 8th Southern African Base Metals Conference 6–8 July 2015, Zambezi Sun Hotel, Victoria Falls, Livingstone, Zambia N SCHOOL Production of Clean Steel 13–14 July 2015, Emperors Palace, Johannesburg N CONFERENCE Virtual Reality and spatial information applications in the mining industry Conference 2015 15–17 July 2015, University of Pretoria, Pretoria N CONFERENCE MINPROC 2015: Southern African Mineral Beneficiation and Metallurgy Conference 6–7 August 2015, Vineyard Hotel, Newlands, Cape Town N CONFERENCE The Danie Krige Geostatistical Conference 2015 19–20 August 2015, Crown Plaza, Johannesburg N CONFERENCE MINESafe 2015—Sustaining Zero Harm: Technical Conference and Industry day 26–28 August 2015, Emperors Palace Hotel Casino, Convention Resort, Johannesburg N CONFERENCE Formability, microstructure and texture in metal alloys Conference 2015 15–17 September 2015 N CONFERENCE World Gold Conference 2015 28 September–2 October 2015, Misty Hills Country Hotel and Conference Centre, Cradle of Humankind, Muldersdrift N 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 N COLLOQUIUM 13th Annual Southern African Student Colloquim 2015 20 October 2015, Mintek, Randburg, Johannesburg N CONFERENCE Young Professionals 2015 Conference 21–22 October 2015, Mintek, Randburg, Johannesburg N CONFERENCE AMI: Nuclear Materials Development Network Conference 28–30 October 2015, Nelson Mandela Metropolitan University, North Campus Conference Centre, Port Elizabeth N SYMPOSIUM MPES 2015: Twenty Third International Symposium on Mine Planning & Equipment Selection 8–13 November 2015, Sandton Convention Centre, Johannesburg, South Africa

INTERNATIONAL ACTIVITIES 2015 12–13 May 2015 — Mining, Environment and Society Conference: Beyond sustainability—Building resilience Mintek, Randburg, 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 14–17 June 2015 — European Metallurgical Conference Dusseldorf, Germany, Website: http://www.emc.gdmb.de 14–17 June 2015 — Lead Zinc Symposium 2015 Dusseldorf, Germany, Website: http://www.pbzn.gdmb.de 16–20 June 2015 — International Trade Fair for Metallurgical Technology 2015 Dusseldorf, Germany Website: http://www.metec-tradefair.com 6–8 July 2015 — Copper Cobalt Africa Incorporating The 8th Southern African Base Metals Conference Zambezi Sun Hotel, Victoria Falls, Livingstone, Zambia 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 13–14 July 2015 — School Production of Clean Steel Emperors Palace, 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

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Fax: +27 11 838-5923/833-8156 E-mail: raymond@saimm.co.za Website: http://www.saimm.co.za 19–20 August 2015 — The Danie Krige Geostatistical Conference: Geostatistical geovalue —rewards and returns for spatial modelling Crown Plaza, 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 25–27 August 2015 — Coal Processing – Unlocking Southern Africa’s Coal Potential Graceland Hotel Casino and Country Club Secunda Contact: Ann Robertson Tel: +27 11 433-0063 26–28 August 2015 — MINESafe 2015—Sustaining Zero Harm: Technical Conference and Industry day Emperors Palace Hotel Casino, Convention Resort, 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 15–17 September 2015 — Formability, microstructure and texture in metal alloys Conference 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

15–17 July 2015 — Virtual Reality and spatial information applications in the mining industry Conference 2015 University of Pretoria 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 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.z

6–7 August 2015 — MINPROC 2015: Southern African Mineral Beneficiation and Metallurgy Conferene Vineyard Hotel, Newlands, Cape Town Contact: Raymond van der Berg Tel: +27 11 834-1273/7

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

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INTERNATIONAL ACTIVITIES 2015 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

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

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

17–18 May 2016 — The SAMREC/SAMVAL Companion Volume Conference 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

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

8–13 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 Conference 2016 Botswana 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

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17–20 July 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 16–19 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

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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

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

Company Affiliates The following organizations have been admitted to the Institute as Company Affiliates AECOM SA (Pty) Ltd

Engineering and Project Company Ltd

Namakwa Sands (Pty) Ltd

AEL Mining Services Limited

eThekwini Municipality

New Concept Mining (Pty) Limited

Air Liquide (PTY) Ltd

Evraz Highveld Steel and Vanadium Corp Ltd

Northam Platinum Ltd - Zondereinde

AMEC Mining and Metals AMIRA International Africa (Pty) Ltd

Exxaro Coal (Pty) Ltd

Outotec (RSA) (Proprietary) Limited

Exxaro Resources Limited

ANDRITZ Delkor(Pty) Ltd

PANalytical (Pty) Ltd

Fasken Martineau

Anglo Platinum Management Services (Pty) Ltd

Osborn Engineered Products SA (Pty) Ltd

Paterson and Cooke Consulting Engineers (Pty) Ltd

FLSmidth Minerals (Pty) Ltd

Anglo Operations Ltd

Fluor Daniel SA (Pty) Ltd

Anglogold Ashanti Ltd

Franki Africa (Pty) Ltd Johannesburg

Polysius A Division Of Thyssenkrupp Industrial Solutions (Pty) Ltd

Atlas Copco Holdings South Africa (Pty) Limited

Fraser Alexander Group

Precious Metals Refiners

Glencore

Rand Refinery Limited

Aurecon South Africa (Pty) Ltd

Goba (Pty) Ltd

Redpath Mining (South Africa) (Pty) Ltd

Aveng Moolmans (Pty) Ltd

Hall Core Drilling (Pty) Ltd

Rosond (Pty) Ltd

Axis House (Pty) Ltd

Hatch (Pty) Ltd

Royal Bafokeng Platinum

Bafokeng Rasimone Platinum Mine

Herrenknecht AG

Roymec Tecvhnologies (Pty) Ltd

Barloworld Equipment -Mining

HPE Hydro Power Equipment (Pty) Ltd

RSV Misym Engineering Services (Pty) Ltd

BASF Holdings SA (Pty) Ltd

Impala Platinum Limited

Rustenburg Platinum Mines Limited

Bateman Minerals and Metals (Pty) Ltd

IMS Engineering (Pty) Ltd

SAIEG

BCL Limited

JENNMAR South Africa

Salene Mining (Pty) Ltd

Becker Mining (Pty) Ltd

Joy Global Inc. (Africa)

BedRock Mining Support (Pty) Ltd

Leco Africa (Pty) Limited

Sandvik Mining and Construction Delmas (Pty) Ltd

Bell Equipment Company (Pty) Ltd

Longyear South Africa (Pty) Ltd

BHP Billiton Energy Coal SA Ltd

Lonmin Plc

Blue Cube Systems (Pty) Ltd

Ludowici Africa

Bluhm Burton Engineering (Pty) Ltd Blyvooruitzicht Gold Mining Company Ltd

Lull Storm Trading (PTY)Ltd T/A Wekaba Engineering

BSC Resources

Magnetech (Pty) Ltd

Sebilo Resources (Pty) Ltd

CAE Mining (Pty) Limited

Magotteaux(PTY) LTD

SENET

Caledonia Mining Corporation

MBE Minerals SA Pty Ltd

Senmin International (Pty) Ltd

CDM Group

MCC Contracts (Pty) Ltd

Shaft Sinkers (Pty) Limited

CGG Services SA

MDM Technical Africa (Pty) Ltd

Sibanye Gold (Pty) Ltd

Chamber of Mines

Metalock Industrial Services Africa (Pty)Ltd

Smec SA

Concor Mining

Metorex Limited

SMS Siemag South Africa (Pty) Ltd

Concor Technicrete

Metso Minerals (South Africa) (Pty) Ltd

SNC Lavalin (Pty) Ltd

Council for Geoscience Library

Minerals Operations Executive (Pty) Ltd

Sound Mining Solutions (Pty) Ltd

CSIR-Natural Resources and the Environment

MineRP Holding (Pty) Ltd

SRK Consulting SA (Pty) Ltd

Mintek

Time Mining and Processing (Pty) Ltd

Department of Water Affairs and Forestry

MIP Process Technologies

Tomra Sorting Solutions Mining (Pty) Ltd

Deutsche Securities (Pty) Ltd

Modular Mining Systems Africa (Pty) Ltd

TWP Projects (Pty) Ltd

Digby Wells and Associates

Runge Pincock Minarco Limited

Ukwazi Mining Solutions (Pty) Ltd

Downer EDI Mining

MSA Group (Pty) Ltd

Umgeni Water

DRA Mineral Projects (Pty) Ltd

Multotec (Pty) Ltd

VBKOM Consulting Engineers

Duraset

Murray and Roberts Cementation

Webber Wentzel

Elbroc Mining Products (Pty) Ltd

Nalco Africa (Pty) Ltd

Weir Minerals Africa

L

xiv

APRIL 2015

Sandvik Mining and Construction RSA(Pty) Ltd SANIRE Sasol Mining(Pty) Ltd Scanmin Africa (Pty) Ltd

The Journal of The Southern African Institute of Mining and Metallurgy

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