MASTER'S THESIS
Challenges with Deep Mining Case Studies at Mindola and Mufulira Underground Mines, Zambia
Kristoffer Berglund Daniel Günther 2014
Master of Science (120 credits) Natural Resources Engineering
Luleå University of Technology Department of Civil, Environmental and Natural Resources Engineering Challenges with deep mining
Case studies at Mindola and Mufulira underground mines, Zambia
Kristoffer Berglund & Daniel Günther
MSc-thesis in Rock Mechanics and Rock Engineering
Department of Civil, Environmental and Natural Resources Engineering
Lulea University of Technology
SE – 971 87 Lulea
Sweden
ii
Preface
This MSc-thesis in Rock Mechanics and Rock Engineering is the final work that finishes our studies towards an MSc-degree in Natural Resources Engineering. It stands for 30Hp and was carried out during the spring and summer of 2013 at the Department of Civil, Environmental and Natural Resources Engineering at Lulea University of Technology.
First of all we would like to thank Prof. Erling Nordlund and Dr. Jenny Greberg at LTU for granting us the opportunity to conduct this thesis in collaboration with the I2Mine project. This gave us the chance of a life time to visit the mines in Zambia and experience and learn things we never would have learnt elsewhere. This has been incredibly inspiring and we are forever grateful.
We would also like to extend our great appreciation to the staff of the Mufulira and Mindola mines; for their great hospitality, the time and the experience they shared with us to make this study possible.
We would also like to thank Dr. Ping Zang and Dr. Daniel Johansson for their guidance and the inspiring discussions we have shared.
Finally we would like to thank PhD. Student Sraj Banda who arranged the trip to Zambia and took care of all the practicalities associated to the mine visits. We would also like to thank him for the inspiration, help, encouragement, advice and guidance he has given us during this study, without you it would not have been possible.
Personal acknowledgements by the authors:
First of all I would like to thank my friend Kristoffer Berglund for all the good times we have shared during our studies. I would like to thank my father for all the proofreading you have had to endure and for all the guidance you have given me through my entire study time you have been a huge support and I am forever grateful. I would also like to thank my mother for always believing in me providing me the support and confidence to reach goals I did not think I could reach. My fiancée Ida the joy in my life who has always stood by my side, words cannot express what you mean for me. Last I would like to dedicate all my work in this report to my beloved grandpa who passed away during my studies in April 2011.
Daniel Günther Luleå September 2013
iii
I would like to thank my family for always being there for me during my time as a student. A special thanks to my friend Daniel Günther for all the fun and mischief we have experienced together.
Kristoffer Berglund Luleå September 2013
iv
Abstract In pace with our ever increasing demand for metals and with many of the known deposits already exploited, the mining industry is forced to enter into unexplored territories in the search for new deposits. The depth of our earth is one of them; mining operations are currently taking place at greater depths than ever before and exploration into the depths continues. But with the depth new challenges to mining has risen and existing challenges has become even greater. This study will try to shine some light on these challenges, their possible solutions and the considerations that has to be made regarding them.
This report is divided into two parts; part A and part B. In part A, two case studies are presented from two deep mines in Zambia. These mines were visited and investigated for this study. In part B the challenges these mines face due to the depth of mining are presented and analyzed. The analysis is based on literature studies, information gathered on site, basic calculations and numerical modeling performed in the finite element program Phase2 by RocklabTM.
The case studies were performed as a part of the I2Mine research project. The I2Mine research project aims to realize the vision of an invisible, zero impact deep mine. The mines that were studied are Mufulira Copper mine and Mindola Copper mine which are located to the Copperbelt region of Zambia. Both are owned and operated by Mopani Copper Mines Plc. which is the main employer in the region.
The Mufulira mine has a proven and probable ore reserve of 19 million Tonnes and has at the current copper price an expected mine life of 25 years. The main mining method in Mufulira is a variant of Sub level Open Stoping locally referred to as Mechanized Continuous Retreat. The mine is made out of three ore bodies which dip at around 45°, the ore is extracted in stopes and no backfilling is employed since the mining progresses downward. Current mining level is at a depth of 1540 meters and both ore and waste rock is hoisted to the surface. Ventilation is performed mechanically with no complementing air-conditioning. The rock mass failures are generally stress induced and the mine also experience some mining induced seismicity mostly connected to pillar burst and caving of the stopes.
The Mindola mine is a copper and cobalt mine which has a total measured and indicated ore reserve of 28 million tones and an annual production of 1.1million tones putting the life of mine at 25 years. The mining method used for production at Mindola Shaft is a variation of sublevel stoping known as the Vertical Crater Retreat (VCR) method. The ore body in Mindola is on average 12 meters wide with dip varying from 56° to 72°, the ore is extracted in stopes and backfilling is employed. The current production level is located at 1570 m depth and plans are made to deepen the mine further in order to extend its life. The ventilation system in the Mindola mine is based on a technique that creates a pressure difference between the mine and the surface by sucking air out of the mine and letting fresh air in. The most common failure mechanisms
v experienced are slow and time dependent deformations and in some cases stress concentrations in combination with unfavorable rock structures.
After conducting the two case studies at Mindola and Mufulira it was evident that Mufulira was experiencing great geotechnical challenges due to stress induced by the mining method. The main problem is that the MCR method leaves behind crown pillars which acts like stress windows and causes high stresses in the footwall drifts. In Mindola they have faced a similar problem but they dealt with it by changing the mining method, so that they now backfill the stopes. By that they don't have to leave behind any crown pillars which would have acted as stress windows. The change of mining method have also had positive effects on the production as it means less development and that the waste no longer has to be hoisted to the surface. By implementing the same changes in Mufulira similar results could be expected.
A numerical analysis was performed with the objective to investigate how a change in mining method would affect the high stress levels in Mufulira. The numerical analysis was performed using the finite element software Phase2 based on data received from the mine. One of the major goals of the analysis was to determine what a change in mining method would imply for the location of the footwall drifts since these had been moved further and further away from the stopes as the depth had increased.
Results from the numerical analysis indicated that changing the mining method in Mufulira would severely decrease the stress levels in the footwall. The major reason for that was that the change of method would eliminate the stress windows and by that lower the stress in the footwall. This would then mean that the footwall drifts could be placed much closer to the stope, reducing both hauling distance as well as development time.
From the two case studies it could also be concluded that both mines faces great ventilation challenges with respect to heat. The mines reported temperatures in the range of 32°C wetbulb in the production areas at the current mining level. The reason for the high temperatures is that air is heated by both autocompression, which is a function of surface temperature and depth, and by thermal radiation from the host rock, which has at deepest part of the mine an internal temperature of around 50°C due to the geothermal gradient. In order to lower these temperatures the mine plans to expand the vertical infrastructure of the ventilation system and thereby increase the airflow.
With respect to these problems analyses were conducted with the objective to validate the mines proposed solutions and determine if other solutions could be applicable. In the analysis different means of air-conditioning were studied. Two factors of major importance for ventilation with regards to heat were identified: The critical depth and the critical rock temperature. These factors represents the limits were the air basically no longer has any cooling potential due to either autocompression and / or thermal radiation from the rock.
vi
The results indicated that both mines have passed both the critical depth and the depth of the critical rock temperature. The conclusion could be made; that the solution of increasing the air velocity in order to lower the temperatures would only add to the heat load. The recommended solution would therefore be to install air-refrigeration before the air-velocity is increased.
vii
Sammanfattning I takt med vår ständigt ökande efterfrågan på metaller och med många av de kända mineralfyndigheterna redan exploaterade har gruvindustrin tvingats träda in på outforskade områden i sökandet efter nya fyndigheter. Stora djup är ett av dessa områden, gruvdrift pågår för närvarande på större djup än någonsin tidigare och prospektering görs för än större djup, men med ökat djup ökar också utmaningarna. Det är dessa utmaningar och de hänsynstaganden de innebär som denna studie kommer försöka att kasta ljus på.
Rapporten är uppdelad i två delar; del A och del B. I del A, presenteras fallstudier av två djupa gruvor i Zambia, dessa sammanställdes utifrån ett besök i maj 2013 och ligger till grund för analyserna i denna rapport. I del B presenteras och analyseras de utmaningar dessa gruvor står inför med avseende på djupet. Analysen grundar sig på litteraturstudier, insamlad information, grundläggande beräkningar och numerisk modellering utfört idet finita elementprogrammet Phase2 av RocklabTM.
Fallstudierna utfördes i samarbete med forskningsprojektet I2Mine. Forskningsprojektet I2Mine syftar till att förverkliga visionen om en djup osynlig och nollpåverkande gruvdrift. Gruvorna som studerades är två koppargruvor, den ena i Mufulira och den andra i Mindola vilka båda ligger i en region som kallas Kopparbältet i Zambia. Båda gruvorna ägs och drivs av Mopani Copper Mines Plc. som är den största arbetsgivaren i regionen.
Koppargruvan i Mufulira har en bevisad och sannolik malmreserv på 19 miljoner ton och har med dagens metallpris en livslängd på 25 år. Den huvudsakliga brytningsmetoden i Mufulira är en variant av Skivpalls brytning vilken lokalt benämns som Mechanized Continuous Retreat method (MCR). Gruvan utgörs av tre malmkroppar som har en lutning på cirka 45°, malmen utvinns genom att pallar tas ut ur malmkroppen men detta görs utan återfyllning då brytningen fortskrider nedåt. Aktuell gruvdriftsnivå är på ett djup av 1540 meter och både malm och gråberg hissas till ytan. Ventilationen sker mekaniskt utan kompletterande luftkonditionering. De vanligaste brottmekanismerna har i allmänhet varit spänningsinducerade och man har även upplevt några fall av brytnings inducerad seismicitet, dessa har kunnat härledas till pelarbrott samt till större utfall relaterat till produktionen.
Mindola är en koppar och kobolt gruva som har en total känd och indikerad malmreserv på 28 miljoner ton och en årlig produktion på 1.1 miljoner ton malm vilket ger gruvan ett liv på upp till 25 år. Brytningsmetod som används för produktion vid Mindola är en variant av skivpallsbryning kallad Vertical Crater Retreat (VCR)-metoden. Malmkroppen i Mindola är i genomsnitt 12 m bred och dess lutning varierar från 56 till 72 grader. Malmen utvinns genom att pallar tas ut som återfylls med gråberg. Den nuvarande produktionsnivån ligger på 1570 m djup och det finns planer för att fördjupa gruvan ytterligare för att förlänga dess livslängd. Ventilationssystemet i Mindola bygger på en teknik som skapar en tryckskillnad mellan gruvan och ytan genom att suga luft ut ur
viii gruvan och låta frisk luft flöda in. De vanligaste brottmekanismerna som är långsamma och tidsberoende deformationer och i vissa fall spänningskoncentrationer i kombination med ogynnsamma bergstrukturer.
Efter att ha utfört de två fallstudier vid Mindola och Mufulira så var det uppenbart att Mufulira upplever geotekniska utmaningar inducerade av brytningsmetoden. MCR metoden som används i Mufulira lämnar kvar kronpelare som bildar spänningsfönster och som i in tur orsakar höga spänningar i liggväggens orter. I Mindola har de stött på liknande problem, men det löstes genom att ändra brytningsmetod. Mindola ändrade sin brytningsmetod så att de nu återfyller sina stopes, därmed lämnar de inga kronpelare som skulle kunnat agera som spänningsfönster. Detta resulterade i positiva resultat för produktionen då gråberget inte längre transporteras till ytan. Genom att implementera samma förändringar i Mufulira skulle liknande resultat kunna väntas.
En numerisk analys utfördes i syfte att undersöka hur en förändring av brytningsmetod skulle påverka de höga spänningsnivåerna i Mufulira. Den numeriska analysen utfördes genom användning av den finita element program Phase2 och baserades på indata som mottagits från gruvan. Ett av de viktigaste målen med analysen var att bestämma vad en förändring av brytningsmetod skulle innebära för lokaliseringen av fotväggs orterna, eftersom dessa hade flyttats längre och längre bort från malmkroppen i takt med att djupet hade ökat .
Resultatet från de numeriska analyserna indikerade att ett byte av brytningsmetod i Mufulira avsevärt skulle minska spänningsnivåerna i liggväggen. De spänningsfönster som finns idag skulle inte längre vara närvarande vilket skulle reducera spänningen i liggväggen. Därmed skulle det vara möjligt att placera liggväggs orterna närmare malmkroppen vilket skulle minska både transport tid samt tillrednings tid.
Från de två fallstudierna kunde också slutsatsen dras att de båda gruvorna står inför stora ventilationsutmaningar avseende arbetet att kontrollera den stigande temperaturen. Båda gruvorna rapporterade temperaturer i närheten av 32 °C Wet bulb i produktionsområdena på den nuvarande brytningsnivån. Anledningen till de höga temperaturerna är att luften värms upp av autokompression, vilket är en funktion av ytans temperatur och djup , samt genom termisk strålning från det omgivande berget , som har i den djupaste delen av gruvan en inre temperatur på omkring 50 °C på grund av den geotermiska gradienten . För att sänka dessa temperaturer planerar gruvan att expandera den vertikala infrastrukturen av ventilationssystemet och därmed öka luftflödet.
Med fokus på dessa problem utfördes analyser i syfte att validera gruvornas föreslagna lösningar och bestämma om andra lösningar skulle kunna tillämpas. I analysen studerades olika medel för luftkonditionering. Två faktorer som har stor betydelse för ventilation när det gäller värme identifierades: Det kritiska djupet och den kritiska bergtemperaturen. Dessa faktorer utgör de gränser var luften i princip inte längre har någon kylningspotential på grund av antingen autokompression
ix och/eller termisk strålning från berget. All överflödig luftrörelse nedanför detta djup bidrar bara till att öka värmebelastningen på den mänskliga kroppen
Resultaten indikerade att båda gruvorna har passerat både det kritiska djupet och djupet av den kritiska bergtemperaturen. Slutsatsen blev således; den av gruvorna förslagna lösningen med att öka lufthastigheten för att sänka temperaturerna endast skulle få motsatt effekt och leda till en ökad värmebelastning. Den rekommenderade lösningen skulle därför vara att installera ett luftkylningssystem innan lufthastigheten ökades.
x
Glossary
Aquifer A water-bearing bed of porous rock, often sandstone.
Autocompression A term which denotes the fact that the air going underground heats up merely by virtue of the conversion of the ‘potential’ energy of the air at surface into heat when the air moves closer to the center of the earth.
Azimuth A surveying term that references the angle measured clockwise from any meridian (the established line of reference). The bearing is used to designate direction. The bearing of a line is the acute horizontal angle between the meridian and the line.
Backfill Mine waste or rock used to support the roof after ore removal.
Borehole Any deep or long drill-hole, usually associated with a diamond drill.
Competent rock Rock which, because of its physical and geological characteristics, is capable of sustaining openings without any structural support except pillars and walls left during mining (stalls, light props, and roof bolts are not considered structural support).
Contact The place or surface where two different kinds of rocks meet. Applies to sedimentary rocks, as the contact between a limestone and a sandstone, for example, and to metamorphic rocks; and it is especially applicable between igneous intrusions and their walls.
Core sample A cylinder sample generally 1-5" in diameter drilled out of an area to determine the geologic and chemical analysis of the overburden and ore.
Crosscut A passageway driven between the entry and its parallel air course or air courses for ventilation purposes. Also, a tunnel driven from one seam to another through or across the intervening measures; sometimes called "crosscut tunnel", or "breakthrough". In vein mining, an entry perpendicular to the vein.
Crusher A machine for crushing rock or other materials. Among the various types of crushers are the ball mill, gyratory crusher, Handsel mill, hammer mill, jaw crusher, rod mill, rolls, stamp mill, and tube mill.
Deposit Mineral deposit or ore deposit is used to designate a natural occurrence of a useful mineral, or an ore, in sufficient extent and degree of concentration to invite exploitation.
xi
Depth The word alone generally denotes vertical depth below the surface. In the case of incline shafts and boreholes it may mean the distance reached from the beginning of the shaft or hole, the borehole depth, or the inclined depth.
Dew point The temperature below which the water vapor in a volume of humid air at a given constant barometric pressure will condense into liquid water at the same rate at which it evaporates.
Draw level Level where ore can be loaded and removed. A draw level is located beneath the stoping area, and gravity flow transfers the ore to the loading level.
Drift A horizontal passage underground. A drift follows the vein, as distinguished from a crosscut that intersects it, or a level or gallery, which may do either.
Drill A machine utilizing rotation, percussion (hammering), or a combination of both to make holes. If the hole is much over 0.4m in diameter, the machine is called a borer.
Drilling The use of such a machine to create holes for exploration or for loading with explosives.
Dry bulb The temperature in common use referred to in weather reports.
Fault zone A fault, instead of being a single clean fracture, may be a zone hundreds or thousands of feet wide.
Fault A slip-surface between two portions of the earth's surface that have moved relative to each other. A fault is a failure surface and is evidence of severe earth stresses.
Footwall Wall or rock under the ore deposit.
Formation Any assemblage of rocks which have some character in common, whether of origin, age, or composition. Often, the word is loosely used to indicate anything that has been formed or brought into its present shape.
Fracture A general term to include any kind of discontinuity in a body of rock if produced by mechanical failure, whether by shear stress or tensile stress. Fractures include faults, shears, joints, and planes of fracture cleavage.
xii
Geothermal Gradient The almost uniform rate of which the rock temperature rises with depth in a given locale and rock formation. Units of measurement are °F/100 ft or °C/100m. The temperature is strongly associated with heat flow and depends much on the thermal conductivity of the rock as well as the closeness to ingenious activities.
Hanging Wall Wall or rock above an ore deposit.
Haulage The horizontal transport of ore, supplies, and waste. The vertical transport of the same is called hoisting.
Hoisting The vertical transport ore or material.
Incline Any entry to a mine that is not vertical (shaft) or horizontal (adit). Often incline is reserved for those entries that are too steep for a belt conveyor (+17 degrees - 18 degrees), in which case a hoist and guide rails are employed. A belt conveyor incline is termed a slope. Alt: Secondary inclined opening, driven upward to connect levels, sometimes on the dip of a deposit; also called "inclined shaft".
Layout The design or pattern of the main roadways and workings. The proper layout of mine workings is the responsibility of the manager aided by the planning department.
Level System of horizontal underground workings connected to the shaft. A level forms the basis for excavation of the ore above or below.
Lithology The character of a rock described in terms of its structure, color, mineral composition, grain size, and arrangement of its component parts; all those visible features that in the aggregate impart individuality of the rock.
Ore Mineral deposit that can be worked at a profit under existing economic conditions.
Ore Pass Vertical or inclined underground opening through which ore is transferred.
Pillar An area of ore left to support the overlying strata in a mine; sometimes left permanently to support surface structures.
Raise Underground opening driven upward from one level to a higher level or to the surface; a raise may be either vertical or inclined
xiii
Ramp Inclined underground opening that connects levels or production areas; ramps are inclined to allow the passage of motorized vehicles. Ramps usually are driven downward.
Scaling Removal of loose rock from the roof or walls. This work is dangerous and a long bar (called a scaling bar)is often used.
Shaft A primary vertical or non-vertical opening through mine strata used for ventilation or drainage and/or for hoisting of personnel or materials; connects the surface with underground workings.
Stope Underground excavation made by removing ore from surrounding rock.
Strike Main horizontal course or direction of a mineral deposit.
Sublevel System of horizontal underground workings; normally, sublevels are used only within stoping areas where they are required for ore production.
Support The all-important function of keeping the mine workings open. As a verb, it refers to this function; as a noun it refers to all the equipment and materials-- timber, roof bolts, concrete, steel, etc.--that are used to carry out this function.
Waste Barren rock or rock of too low a grade to be mined economically.
Waste Rock That rock or mineral which must be removed from a mine to keep the mining scheme practical, but which has no value.
Wet Bulb The temperature at which water evaporates into the air (at a particular dry bulb temperature) once equilibrium between the water and air has occurred.
xiv
Table of contents 1 Introduction ...... 1 1.1 Background ...... 1 1.2 Scope ...... 1 1.3 Limitations ...... 2 Part A ...... 3 2 General description of Mufulira mine ...... 4 2.1 Geology ...... 6 2.1.1 The ore bodies ...... 8 2.2 Geotechnical conditions ...... 9 2.3 Mining method description ...... 10 2.3.1 Mechanized continuous retreat 1 (MCR 1) ...... 11 2.3.2 Mechanized Continuous Retreat 2 (MCR 2) ...... 12 2.3.3 Development ...... 12 2.4 Mine layout and infrastructure ...... 14 2.4.1 Layout and dimensions ...... 14 2.5 Hoisting system ...... 16 2.5.1 Pillar, Upper and Central sections ...... 16 2.5.2 The Deeps ...... 16 2.6 Ventilation ...... 16 2.7 Mine Planning ...... 17 2.8 Geotechnicalconsiderations ...... 17 2.8.1 Failure ...... 17 2.8.2 Mining induced seismicity ...... 18 2.9 Ground control practice...... 19 2.9.1 Damage mapping ...... 19 2.9.2 Numerical analysis ...... 21 2.9.3 Support standard ...... 21 2.9.4 Rehabilitation ...... 22 2.10 Challenges ...... 22
xv
2.11 The Future ...... 23 3 General description of Mindola mine ...... 24 3.1 Geology ...... 26 3.1.1 Hydrogeology ...... 27 3.1.2 Exploration...... 27 3.2 Geotechnical conditions ...... 29 3.2.1 Rock mass data for mine design ...... 29 3.2.2 Ground behavior ...... 30 3.3 Mining method ...... 32 3.3.1 Main level development ...... 33 3.3.2 VCR sublevel development ...... 33 3.3.3 Charging and blasting ...... 36 3.4 Mine Planning ...... 38 3.4.1 Production planning: ...... 38 3.4.2 Planning challenges ...... 39 3.5 Mine layout and infrastructure ...... 39 3.5.1 Hoisting system ...... 41 3.6 Ventilation ...... 42 3.7 Geotechnical considerations ...... 43 3.7.1 Failure ...... 43 3.7.2 Seismicity ...... 43 3.8 Ground control practice...... 45 3.8.1 Damage mapping ...... 45 3.8.2 Numerical analysis ...... 45 3.8.3 Support standards ...... 45 3.8.4 Rehabilitation ...... 47 3.9 Challenges ...... 47 Part B ...... 49 4 Stress ...... 50 4.1 Numerical Analysis ...... 50
xvi
4.1.1 Model setup ...... 52 4.1.2 Crown pillar ...... 53 4.1.3 Backfill ...... 53 4.1.4 Footwall drive distance from stope ...... 53 4.2 Results ...... 53 4.2.1 Crown pillar ...... 53 4.2.2 Backfill ...... 55 4.2.3 Footwall drive distance from stope ...... 56 4.3 Conclusion ...... 57 5 Ventilation ...... 58 5.1 Heat ...... 58 5.1.1 Autocompression ...... 58 5.1.2 Critical depth and critical rock temperature ...... 59 5.1.3 Heat sources ...... 60 5.1.4 Physiological heat effects on humans ...... 62 5.2 Ventilation techniques ...... 63 5.2.1 Mechanical ventilation ...... 63 5.2.2 Air conditioning ...... 64 5.3 Identified problems ...... 65 5.3.1 Solutions for the near future ...... 66 5.4 Analysis ...... 67 5.4.1 The critical depth due to autocompression ...... 67 5.4.2 The depth of the critical rock temperature ...... 67 5.5 Result ...... 67 5.5.1 Calculations of the critical depth due to autocompression ...... 68 5.5.2 Calculations of the critical rock temperature ...... 68 5.6 Conclusions ...... 69 5.6.1 General conclusions...... 69 5.6.2 Conclusions for Mufulira and Mindola ...... 69 5.7 References ...... 71
xvii
5.7.1 Literature ...... 71 5.7.2 Publications and reports ...... 71 5.7.3 Electronic documents ...... 71 5.7.4 Unpublished material ...... 72 5.7.5 Interviews ...... 72 Appendix A ...... a Appendix B ...... m Appendix C ...... u Appendix D ...... w Appendix E ...... ä Appendix F ...... dd
xviii
1 Introduction
1.1 Background The recent year’s growth of developing countries has led to an increase in the global demands for minerals and metals this has in turn led to higher metal prices than ever seen before and by that opened up for mining industry to explore its horizons. Fictional news such as space mining has figured in renowned newspapers and serious attempts to extract deep sea minerals have been commenced. In the shadow of this, exploration has also continued below the already known shallow deposits deep into the earth’s crust.
As unexploited areas become more and more scarce, all known shallow deposits are soon exploited and land areas becoming more expensive it is likely that most mining operations in the future will be located at great depths. Already today the deepest mines in the world are mining at depths of around 5 kilometers. But with great depths comes great challenges; among of which heat along with rock stress will be the limiting factors to which depths mining operations ultimately can be extended to.
This thesis was performed as a part of work package 2 and 3 in the I2Mine research project. The main focus of the I2Mine project is on developing technologies adapted for deep mining activities and the aims of the project are to realize the vision of an invisible, zero impact mine. Work package 2 focuses on novel mining methods and processes for deep steeply dipping orebodies, and work package 3 focuses on inventory of ground control practices at great depth worldwide for steeply dipping orebodies. The financing for this maters thesis was provided by the European Commission 7th Framework Programme Objective
The objective with this master thesis is to perform case studies of two deep mines and through these investigate and identify the challenges these mines face with regards to the depth. Based on the results from these studies the challenges will further be analyzed with the goal to identify solutions which will minimize their current effect on the mining operations.
1.2 Scope This Master Thesis involved two case studies that were performed on site at Mindola and Mufulira mines, Zambia. The case studies were conducted through interviews with different members from the mine operations staff, gathering of unpublished material at the mine and through our own observations. A comprehensive literature review on mine ventilation and its applications along with a literature study on mining methods were performed. The identified challenges were then analyzed through calculations and numerical modeling in the finite element program Phase2 by RocklabTM.
1
1.3 Limitations This study focuses on the challenges with regards to the depth that the two mines face. These are described in Part B of this report. All the other challenges that the mines face are considered to be out of the scope of this study and have therefore been left out in the analysis. The analyses that were performed have only been based on the information that was made available for the authors during the case study. There may therefore exist information that could slightly alter the result and change the conclusions. No considerations have further been taken to economic feasibility of the proposed recommendations these have only been made as a direct solution to the investigated problem.
2
Part A
Two case studies
3
2 General description of Mufulira mine The Mufulira mine site lies approximately 50 km north of the town Kitwe, close to the town Mufulira in the Copperbelt region of Zambia see Figure 1. Mufulira has approximately 140,000 inhabitants (Encyclopedia Britannica academic online edition, 2013) of which a majority works in connection with the mine. The mineralization was first discovered in 1925 and the mine has been in operation since 1933, today it is owned and operated by Mopani Copper mines plc. which is owned by an investment company, Carlisla Investment Corporation (ZCCM, 2013) (Kampeni, 2013).
Mopani employs over 8,000 direct employees with a complement of contract labor of more than 7,800 thereby making the total labor strength of approximately 16,000. Out of this the Mufulira mine employs approximately 5000 employees.
Mining at the Mufulira mine site is today commenced in two underground mines within the site; Mufulira main Central and Mufulira East Portal. Mining operations used to extend over 5.5 km of strike length, but since the Mufulira West deep mining area was mined out and closed down, current operations now take place over a 3 km strike length.
The total ore reserves in the Mufulira mine including the shallow oxide ore is with a cutoff grade of 1.5%; 19 million Tonnes only including proven and probable resources, with the inferred resources included it is; 20 million Tonnes. Out of this the Mufulira main Central stands for 16 million Tonnes. The annual production of the mine is 1.5million Tonnes but after a newly commenced shaft down to the 2020 m level is developed the production is expected to increase to 2 million Tonnes. The mine has with the current metal price and with the reserves down to the 1640 m level a mine life to 2024, but with a new shaft to the 2020 m level that is currently under construction, the life of the mine will be extended to 25 years.
There are three types of mining methods currently in use at the Mufulira mine; In the Mufulira East Portal, the mining method employed is Room and Pillar with waste rock backfill. Mining operations is commenced down to its deepest level of 140m; the ore is then trucked to the surface and treated in the VAT-Leach plant. In the Mufulira main central the mining method is called Mechanized Continuous Retreat 1 and Mechanized Continuous Retreat 2 (MCR1 and MCR2). These are essentially variations of the sublevel open stoping mining method and are the most widely employed methods in the Mufulira mine. In the shallower part of the main central, Room and Pillar with waste rock backfill is employed. As a complement to the mentioned mining methods the mine also employs In-situ leaching of old stopes. Due to the nature of this report the focus of this case study has been on the mining operations in the Mufulira main Central and especially on the Mufulira Deeps section.
The Mufulira main central comprises of eight vertical, one sub vertical and three sub inclined shafts that provide logistical support to the mine from the surface to the Mufulira Deeps section.
4
The Mufulira Deeps section is the deepest area in the Mufulira mine Central and mining is currently commenced down to the 1540 m level. The shift schedule is 8 hours shift 3 times á day so that a constant production of 24 hours á day 365 days per year can be maintained. All the development is made by contractors, the operations at the mine are 80% mechanized and nothing is automated (MOPANI, 2013a).
Figure 1 Location of Mufulira(Google maps, 2013)
5
2.1 Geology The Mufulira ore body lays within the geological formation known as the Neoproterozoic-Lower PalaeozoicLufilian arc which is a segment of the Pan-African orogenic system, see Figure 2. The copperbelt area which is illustrated in Figure 3, is underlain with a basement complex, at Mufulira this complex is named the Lufubu group or locally called Lufubu schist and comprises of a massive micaceous and chloritic quartzite.
Figure 2 The Lufilian arc, the square marks an area in which the Copperbelt lies this area is presented in figure 5. (Wendorff, 2005, s. 62)
6
Figure 3 Local geology in the northern part of the Lufilian arc and the copperbelt, Mufulira can be seen in the Southeast corner. Map symbols: (1) pre-Katangan basement; (2) Roan Group unconformably overlying basement; (3) undifferentiated post-Roan units of the Katanga Supergroup; (4) undifferentiated complexes of fragmental rocks; (5) boundary between fold-thrust belt and equivalent complexes in DRC; (8) location of unconformable lower boundary of Petit Conglomerat in Zambia; (9) selected localities with Fungurme Group foreland succession in the DRC and equivalent complex in Zambia; (10) approximate S limit of Fungurume Group. (Wendorff, 2005, s. 63)
The Lufubu schist is overlain by the base of the Lower Roan Group, the footwall formation. This formation consists of banded and cross-bedded Aeolian and aqueous quartzite which thickness depends on the basement topography and therefore varies.
The footwall formation can be divided in to three parts: The lower-, the middle- and the upper- part. The lower part which overlies the inconsistency between the Basement complex of the Roan group and the Basal conglomerate is believed to be of Aeolian origin and consists of a dark grey and finely bedded quartzite. The middle part is considered to consist of well-bedded quartzite which are white to dark banded and also less clayey than the underlying quartzite. The upper part also consists of a quartzite; this quartzite is the least clayey of the three and is also more calcareous.
7
The footwall formations are succeeded by the Mufulira ore bodies (Wendorff, 2005) (MOPANI, 2013b) (Sarpong, 2013).
2.1.1 The ore bodies The Mufulira mine is comprised of three ore bodies, referred to as orebody A, B and C. Ore body A is the smallest of the three and stretches horizontally 300 m, orebody B is slightly larger and stretches horizontally 600 m, while orebody C is the main ore body, this mineralization is also mined in the mine in Mindola approximately 40 km south of Mufulira. But the ore body itself is limited by the cut off grade of 1.5% to a horizontal stretch of 1400 meters and has been followed down to a depth of at least 2000 m. The ore body generally dips 45 degrees in the North East direction but at some places it flattens towards 35 degrees.
The ‘C’ orebody comprises mineralized grey quartzite, which is present throughout the mine area and is underlain by a mud seam. The ‘B’ orebody comprises mineralized grey sericitc quartzite. The ‘A’ horizon comprises a pink to dark grey feldspathic generally massive quartzite with some greywacke carbonaceous lenses.
The orebody contact is generally distinct in all the ore bodies; In C they have a mud seam, in B they have a clear-cut and the ore and host rock is pink respectively black, in A they also have a clear- cut with the pink quartzite against the grey argelite, but in the hanging wall the contact is not as clear. Table 1 summarizes the different orebody characteristics.
The hangingwall formation overlaying the ore bodies, mainly consists of the argillaceous and dolomites. These sediments indicate an evident change from the pre-dominantly arenaceous rocks of the Lower Roan Group to the Dolomite rocks of the Upper Roan Group (MOPANI, 2013b) (Sarpong, 2013).
Table 1 Orebody characteristics. Orebody A B C Average Grade 5% 3% 2.5%-3% Stretch 300 m 600 m 1400 m Average Dip 45o NE 45o NE (35o -)45o NE Hangingwall geology Strong Argelite Dolomite Mix of Dolomite and Shales
8
2.2 Geotechnical conditions The Mufulira rock mass generally consists of a sequence of strong sedimentary formations overlying competent basement Schist and Granites, with rock strengths that range from 200 to 300 MPa. The Mufulira orebodies generally dips on average 45 degrees to the Northeast, but flatter areas exist were the dip is35 degrees. The three ore bodies are described in; stratigraphic units, thickness and general geotechnical competency in Table 2. The rock mass below 500m level is by Copperbelt standards generally very strong. Due to this the mine experience brittle stress fractures and seismic activity in the areas below 500 m. In the enrichment zone above 500 m the problems concerns the effect of water in the rock mass (MOPANI, 2013b) (Sarpong, 2013).
Table 2 The General Stratigraphy and Geotechnical Characteristics in the Deep Mining Area. (MOPANI, 2013b) Geological Unit Thick UCS RMR General Geotechnical Description (MPa) Lower argillaceous quartzite 40m 320 - Good quality, massive, bedded, partings infilled with Calcite. Good quality, but bedded with weak graphitic partings in lower, greywacke part of orebody. Upper ‘A’ orebody quartzite 8m 220 81 part shows calcite infilled bedding partings with beds 10 to 50cm thick. InterA/B quartzite 320 Banded shales, quartzite and 68 Average quality, Bedded dolomite 10-30m 160-250 Lower dolomite 50 49 Weak and friable ‘B’ orebody quartzite 7m 320 81 Very good quality, bedded. Inter B/C quartzite 310 Good quality, bedded Banded shales& dolomites 4-18m 64 Moderate quality. Mud seam at base may be weak in 211 places. ‘C’ orebody quartzite 9m 310 81 Very good quality Footwall quartzite 0-150m 270 61 Shear zone at basement contact Basement Schist - 300 - Strong competent
9
2.3 Mining method description The mining method in Mufulira is called Mechanized continuous retreat, which essentially is a mining method called sublevel open stoping (SLOS). SLOS is a mining method especially suited for steeply dipping, regular and competent orebodies that have stable rock in both hanging wall and footwall. The ore is recovered through the excavation of large open stopes which are commonly backfilled, see illustration in Figure 4. Both vertical (sill pillars) and horizontal pillars (crown pillars) are usually left to separate the stopes. As the name implies sub levels are created between the main levels for drilling and blasting access to the stope. Beneath the stopes a draw level is created where the Load Haul and Dump (LHD) trucks can muck out the ore for further transportation usually by trucks or carts to the crusher.
Figure 4 General layout of a stope in the Sublevel Open Stoping mining method (Atlas Copco, 2003, s. 23)
The mining method in Mufulira is further divided into two variants; MCR1 and MCR2 with the main difference that the stope height is double in MCR2 compared to MCR1. The following text describes the MCR and the two variations of it (Atlas Copco, 2003) (MOPANI, 2013c).
10
2.3.1 Mechanized continuous retreat 1 (MCR 1) The mining layout of MCR 1, see Figure 5, comprises of sublevels spaced at 17m vertical intervals. At each of these sublevels, development, production, sequential production drilling and ore extraction operations are conducted. The Ore body is accessed by an access drive from the ramp to the footwall drive. The footwall drive is 4.0m x 4.0m and is developed along the strike of the Ore body at a minimum distance of 30m from the geological contact. Ore body cross cuts that also are 4.0m x 4.0m are developed at 50 m intervals along the length of the footwall drive. In the geological footwall contact a mining drive also of 4.0m x 4.0m is developed. Where the ore body is very wide the mining drive is complemented by a second parallel mining drive within the ore body. The mining drive follows the dip of the Ore body and its position is optimized to ensure maximum ore recovery and long hole drilling efficiency. 50 meters apart at the end of each stub cross cuts in the geological hanging wall contact; slot raises of 2.0m x 2.0m are developed. The stub cross cuts are mined to sizes of 3.8m x 3.8m. At every 100 meters interval vent and tip cross cuts are developed, these also incorporate 1.8 m diameter exhaust ventilation raises and 1.8m diameter ore passes (MOPANI, 2013c).
Figure 5 Layout of MCR 1 (MOPANI, 2013c)
11
2.3.2 Mechanized Continuous Retreat 2 (MCR 2) The overall layout of MCR2 is very similar to MCR1 see Figure 6, sublevels are also here developed at 17m vertical intervals but in this mining method a crown pillar is left at the top of every third sublevel. This allows the use of every second level as an extraction level. The first sub level is only used for production drilling and no ore extraction is performed at this level. Due to this slight change in the mining method, the design is just slightly different compared to MCR1: Since no extraction is performed at the drilling levels the crosscuts can be spaced further apart, therefore the crosscuts in MCR2 is placed 50m apart instead of 25m as in MCR1. The crosscuts at the extraction levels have however the same spacing, 25m as in MCR1 but slot raises are placed at 100 m intervals. The closer spacing of the crosscuts at the design levels is to ensure that maximum ore extraction is achieved. The rest of the ancillary infrastructure such as ore passes and ventilation raises are designed in a similar way as in MCR1 (MOPANI, 2013c).
Figure 6 Layout of MCR 2 (MOPANI, 2013c)
2.3.3 Development The development drilling at Mufulira is done both with the use of Jackhammers and by electro hydraulic rigs (Jumbo). The fleet of electro hydraulic rigs comprise of a variety of face rigs from Atlas Copco Ltd, Sandvik mining and Construction Ltd.
12
Rock from the development is transported using LHDs, the standard size of the equipment used is LHD’s with a bucket capacity of 4.6 cubic meters however smaller loaders are also operated. Development requirements are higher in MCR 2 as compared to development in MCR 1.
At the present all the development is made by contractors, but lately the goals for the development have not been reached. Due to the criticality of maintaining the mining sequence and keeping up with the development, the Mine has placed orders to buy their own equipment so that they can take care of the most critical parts of the development themselves (MOPANI, 2013c) (Kampeni, 2013).
2.3.3.1 Vertical development All the drifts, ramps and crosscuts are excavated by the use of conventional drill and blast techniques. Ventilation raises and ore passes are generally excavated through raise boring. The new production shaft down to the 2020 level will be excavated in two phases; the first phase will be the excavation of the shaft from the current depth of the mine to the surface, this part will be raise bored, in the second phase the lower part of the shaft will be excavated from the 1587m level down to the 2020m level through blind sinking. The excavation will commence in September- October 2013 and the shaft is expected to be put in to production during 2016 (Kampeni, 2013).
2.3.3.2 Production drilling Production drilling is performed with the use of conventional drifter machines and longhole electro hydraulic rigs from the mining drives. The holes are drilled in a fan pattern and are directed upwards, generally the ring spacing is 1.8m and the toe burden is 2.1m. The holes are drilled upwards to the bottom of the crown pillar. Production holes are drilled to diameters of 76mm and in rare cases to the diameter of 89mm where reaming is needed. The production drilling in MCR2 is done from the mining drives at the drilling level and the extraction level, the drilling pattern is the same as in MCR1 (Kampeni, 2013).
2.3.3.3 Stoping operations Stoping is initiated by the creation of a slot cut-out across the width of the ore body from the position of the slot raise. The stope is excavated by firing rings positioned in the cross cut using the slot as a breaking face. After completion the slot cut then serves as a breaking face for the longhole blasting. The stoping is always performed in retreat. The blast holes are charged with bulk emulsion explosives. In order to ensure that the stope profile is maintained a slot at an interval of 50 m is developed. Ore extraction from the stopes is performed with LHD’s. The ore is transported through the cross cuts and then it’s either hauled to an ore pass or loaded on to a dump truck depending on the proximity to the ore pass. The stoping procedure is very similar for the two mining methods but in MCR2 the lower sub level lags behind the top sub level by 3 rings approximately 5.4m (MOPANI, 2013c) (Kampeni, 2013).
13
2.3.3.4 Automation There is today no automation in the mine and the mine is only 80% mechanized. There are also no plans for the near future to bring atomization to the mine (Kampeni, 2013).
2.4 Mine layout and infrastructure The mine is made up of three ore bodies referred to as A, B and C where C is the major ore body that have been followed to a minimum depth of 2000 m. The mine has 9 shafts of which 7 are designated to ore and rock handling, see Figure 7. The access to the mine is made through two shafts; one vertical shaft down to 500 m and then through one sub-vertical shaft down to 1400m (Kampeni, 2013).
Figure 7 Ore/Waste flow chart (MOPANI, 2013c)
2.4.1 Layout and dimensions The access to the ore body is made by developing an access drift from the 4.8m x 5.5m decline. At the end of the access cross cut, a 4.0m x 4.0m footwall drive is developed along the strike of the ore body. The footwall drive is positioned 30 m away from the geological footwall contact of the ore body but this distance is being increased to 40-50 m since the mine is experiencing stress problems partly due to greater depths. On the mining level the ore body cross cuts are developed at 50 m intervals along the length of the footwall but on the draw level the cross cuts are spaced at 25m intervals in order to get a higher recovery. The ore body cross cuts as well as the mining drives at
14 the contact of the geological footwall of the ore body is also 4.0m x 4.0m, a summary of the dimensions is presented in Table 3.
Table 3 Dimensions of the different drifts and raises in the mine. Locations Dimensions Ramp 4.8 x 5.5m Footwall drive 4.0 x 4.0m OB Cross cuts 4.0 x 4.0m Mining drive 4.0 x 4.0m Slot raise 2.0 x 2.0m Stub cross cut 3.8 x 3.8m Ventilation raise (old) Ø 1.8m Ventilation raise (new) Ø 2.4m
A slot raise of size 2.0m x 2.0m is developed on the geological hanging wall contact of the ore body at 50m intervals at the end of each stub cross cut. Stub cross cuts are mined to sizes of 3.8m x 3.8m. The position of the mining drive is dictated mainly by the dip of the ore body and is optimized to ensure maximum ore recovery and long hole drilling efficiency. Where an ore body is very wide, two mining drives are placed in the ore body. The layout also includes vent cross cuts and tip cross cuts at 100m intervals. At the end of these cross cuts exhaust ventilation raises and ore passes with a diameter of 1.8m are placed. Due to great ventilation problems in the deeper production areas the vent raises are now being expanded to 2.4 m diameter. In the MCR2 the lower sub level lags behind the top sublevel by 3 rings (5.4m). The stope layout applied for MCR2 is presented in Figure 8 (MOPANI, 2013c).
Figure 8 Stope layout, red = 4 x 4m footwall drifts, blue = 4 x 4m mining drive, green = 4 x 4m ore body crosscuts spaced at 50m intervals on mining level respectively 25m intervals on draw level.
15
2.5 Hoisting system
2.5.1 Pillar, Upper and Central sections Ore from the Pillar, Upper and Central section’s is handled using any of the major tramming levels located at 730m level, 1040m level 1140 m level, 1240m level and 1340 m level. Ore passes connect these levels to the central ore stream that runs directly to the 1360 m level crusher via the Matelo shafts, see Figure 7.
Waste generated from the development is tipped into waste passes and loaded from chute boxes into grandby cars on the major tramming levels and trammed using locomotives to the 56 tip located behind the Musombo sub-vertical shaft. This tip is positioned in the 56 block area and extended from 880 m level to 1340 m level. The rock is then hoisted to 430 m level via the Musombo sub vertical shaft. The skip capacity on Musombo SV is 6 tons. From 430 m level the rock is hoisted out to surface via Kanono shaft (MOPANI, 2013c).
2.5.2 The Deeps After extraction from the stopes, ore from the Deeps section of Mufulira mine is loaded into AD55 caterpillar dump trucks either directly at draw point loading bays or from chute boxes and is then transported to the 1300m level and tipped directly to the 1360 crusher. The ore is then transported via plan ‘C’ conveyor to 1040 m level. The inclination of the plan ‘C’ conveyor is 45 degrees. From 1040 m level the ore is skipped via Matelo 1 and Matelo 2 inclined shafts to the 430m level, the skips has a capacity of 14 tons each and the shafts has a 45 degree inclination. The ore is finally skipped to the surface via the Lubwe shaft which has a skip capacity of 10 tons.
Waste generated from the development which ends in the deeps is also loaded on trucks. The waste is then transported to 1300mL and tipped into a waste tip. This waste rock is loaded into grandby cars from chute boxes and is then trammed on the 1340m level using locomotives to a tip located in block 56 behind the Musombo sub-vertical shaft. The rock is then skipped to 430 m level via the Musombo sub-vertical shaft, the skip capacity is 6 tons. From the 430 m level the rock is then hoisted to the surface via the Kanono shaft (MOPANI, 2013c).
2.6 Ventilation The ventilation system in the Mufulira mine is based on the technique to create a pressure difference. This is done by having up cast fans that sucks out the air from the mine creating a lower pressure which forces the fresh air to seek its way down in to the mine through the access shafts. The ventilation system is built up of exhaust shafts from the main levels to the surface with diameters of 4m and ventilation rises with a diameter of 1.8m between the productions levels spaced at every 100 m. The volume of air circulated in and out of the mine is approximately 1200 m3/s (Mutawa, 2013).
16
2.7 Mine Planning The mine planning can be divided in to two categories; long term planning and short term planning. The long term planning is basically based on geological information, gained through prospect drilling. However geotechnical aspects concerning the design of the footwall development and the mining sequence are also taken into account. This is performed with guidance from the rock mechanics department which for example provides numerical models that determines the optimal production sequence.
The Softwares used for long term planning in Mufulira are the package from GemcomSoftwares™. Gemcoms package includes SURPAC for design purposes and Mine Sched for long term planning and ore reserves generation.
The short term planning is based on the goals in the long term planning, the local geology at the production face and the objective to maintain the mining sequence. One of the greatest challenges in the short term planning is the meandering nature of the ore body. Meaning that both the thickness varies as well as the contact face which often leads to that the geologists has to come down to the stope to map and mark the contact for the drillers which is a time consuming process. The planning is performed with the help of the software Microsoft Excel from Microsoft™, for shift scheduling etc. (Kampeni, 2013).
2.8 Geotechnical considerations
2.8.1 Failure Failures are generally a natural part of a mine and do not have to have negative meaning if they can be predicted and controlled. However in order to predict failures the demand is high for a skilled geotechnical department, especially when mining operations are executed on great depths. (Sarpong, 2013)
2.8.1.1 Failure mechanisms in Mufulira The failure mechanisms experienced in Mufulira are different in different locations of the mine but the failures are in general stress induced: In a limited area in the easternmost parts of the mine, squeezing and depending on how the mine proceeds sometimes also rock bursts has been experienced. In the central part and the western part of the mine the main type of failure mechanism is strain bursts/rock bursts. In the western part of the mine, there is an old mining area where the mining method room and pillar was employed. This area has been re-opened in order to recover the remnant pillars with rock bursts/pillar bursts as an effect.
At certain areas the mine has also been forced to leave pillars behind due to dewatering due to problems with water inflow in the hangingwall. At these areas stress-related problems are experienced. Most of the stress-related problems are located to the crosscuts, stopes or footwall drifts generally with fallouts as a result but sometimes the drifts are squeezed instead.
17
The failures are generally local and the fallout volumes are small but larger collapses has occurred; in 2009 16m of a drift collapsed. This area collapsed due to stressing/de-stressing of the footwall. The area was during mining operation highly stressed and then became de-stressed when the mining had finished. The ground was crushed while it was subjected to the high stresses and when it later was de-stressed the crushed rock fell in to the drifts and parts of the drift even collapsed completely.
Production stop due to fallouts is generally concentrated to the affected stope, drift or crosscut but the size varies so the effect also varies. Fallout has, however, never affected the entire production. There are currently no plans for changing the mining method due to ground control problems, but in the areas with high water inflow rate the mine is forced to use the MCR1 mining method instead of MCR2 (Sarpong, 2013).
2.8.2 Mining induced seismicity Strain bursts are generally the main failure type but at certain parts of the mine seismicity has also been experienced. The seismic problems are mostly located to the western part of the mine and are connected to the pillar bursts in this area. Seismicity has also been experienced when the hanging wall caves into the stope. The reason for this is that the hanging wall is laminated and can therefore form a cantilever with the effect that large rock volumes suddenly can fall out/cave more or less simultaneously. Seismic activity has also been registered when the mining sequence has not been kept. That resulted in great stress concentrations at the stope face since due to the great distance from the faces at other levels. The mine does not have a protocol for the re-entry process. After a seismic event has occurred the area is inspected by listening which is performed by the shift bosses. If the affected area is found to be stable for one shift; the next shift is allowed to enter.
The seismicity in the mine is under constant monitoring with the use of the ISSA system. Figure 9 illustrates the seismic events of magnitude greater than 1 that has been monitored since the system was installed.
The data from the seismic monitoring is used to perform back track analysis and determine what caused the seismic event or failure. The Mine has found that blasting in the stopes often triggered seismic events. The monitoring is part of the environmental program and is used to report to the government what has happened after fallouts. The objective for the distant future is to be able to predict where and when seismic events will occur, but this is yet far away (Sarpong, 2013).
2.8.2.1 Destressing No de-stress blasting is performed. The only preventive de-stressing methods applied is to try and maintain the mining sequence and reduce the amount of pillars left behind as well as reducing the standup time of the crown pillars in the stope.
18
Figure 9 Seismic events monitored in the Mufulira mine (Sarpong, 2013). The size of the ball represents the size of the seismic event 2.9 Ground control practice
2.9.1 Damage mapping Damage mapping is performed three times a week but the same place is never inspected more than once a week. The inspection is carried out by the rock mechanic officers, which each have their own section of the mine. The Mine does not have a certain protocol for damage mapping nor do
19 they document it for the future. The inspections are performed visually were the following is observed, see also Figure 10 and Figure 11:
Loaded mesh and bent face plates Mirror surfaces in the excavations – particularly near/approaching the stope faces – Formation of dog ears in the footwall and mining drives – ahead of the mining fronts or stope faces. Micro-folding and spring-back effects in cross-cuts and mining drives - this could be far away from the stoping areas Pop-up support elements – rebars – or damage steel arched set in burst prone areas
Figure 10 Dog earing in a circular tunnel.
Figure 11 Pop-up support, bolt that has been ejected out of the rockmass due to high stress.
If the damages are observed during the mapping, decision regarding rehabilitation has to be made; these decisions are based on the severity of the damage and the life of the excavation. Usually no
20 rehabilitation is performed, but when it is needed usual remediation technique is to install cable bolts (Sarpong, 2013).
2.9.2 Numerical analysis Numerical analyses are an effective way of predicting failures and determine why failures have occurred. The software used for the analyses in Mufulira are Map 3D, Phase 2 and Flac 3D. The numerical analyses performed in Mufulira are mostly focused on certain mining areas that experience problems. Mine-wide analyses has been performed but is too time consuming due to the lack of computer power. The analyses are therefore used as a tool to determine what kind of geotechnical problem the mine is facing in the problematic areas. This has proven to be very effective and most of the geotechnical problems and the fallouts can be explained through the model. The modeling has also been used to determine the most appropriate mining sequence (Sarpong, 2013).
2.9.3 Support standard The ground control practice is mostly based on traditions but it has been reviewed over the years. Through the reviews a support standard has been developed which involves, bolting, meshing and lacing. The support standard is different for different types of drifts since they have different life length. The support is installed so that its ultimate strength and with the aim to design the life time of the rock support that suits each type of construction. Regular rehabilitation is also common especially in the footwall drifts.
The bolts are mostly fully grouted rebar with the lengths 2.0m and 2.4m but the objective is to standardize the 2.0m bolt. Hydrabolts are also being implemented more frequently; a Hydrabolt is a South African friction bolt that is very similar to the Swellexbolt but less expensive. The main difference between Swellex and Hydrabolts is that the Hydrabolt can contain the water, making it stiffer. The mine also uses tendon straps and mesh. The mesh was chosen instead of shotcrete because it is fast and easy to install and it does not need as much logistics as the shotcrete does. This support is only applied in the footwall drifts and crosscuts and never in the stope because they have a tendency to entangle which disrupts the loading. In the areas with high stresses lacing and bolting is used. This can also be complemented with a finer mesh if the ground is prone to unravel. The support standard can be abandoned in special cases but does always remain as a minimum demand. The maximum unsupported area in the mine is only allowed to be the face +1 m. All the scaling and the support are manually installed by contractors and they have to follow the support standards. The contractors are at all times allowed to put in extra support if they see the need for it. This decision can also be taken by the rock mechanics engineers for that area. The support installation is after installation randomly inspected visually by the mining officers or the mine´s supervisor in rock mechanics and sometimes sampling from the support material is done for testing of the quality in laboratories. If the contractor has not performed the support according to the 21 standard they will not receive any payment for that installation. The following is observed during the inspections:
If enough grout is used (thickness is measured). If the bolts are drilled in faults, If the bolts are loose (pull tested) If the bolts are installed in the correct angle If the bolts cross through the bedding planes. If the mesh and lace are correctly installed.
For some larger excavations such as for workshops or crushers, the support standard is not applicable. For these areas the support installations is based on information such as UCS and joint mapping. The flaw with this approach is that the information is not collected on a regular basis.
Support Standards for MCR1 and MCR2
The support which is applied at Mufulira Mine Site in MCR1 and MCR2 falls within the category of passive-active support. The support recommendation varies from one area of the mine to another depending on the ground conditions. The support standard and roof bolt installation procedure is illustrated in Appendix A (Sarpong, 2013).
2.9.4 Rehabilitation Usually rehabilitation is done after the ground has unraveled. The mine has no apparent system which decision regarding rehabilitation is based on. The rehabilitation is made by separate contractors that rehabilitates under the guidance from the mines own rock mechanics division which inspects the affected area and depending on the life length of the area, decides if or how much rehabilitation is needed.
The mesh is sometimes bled but not often and the rock mechanics do not advise the contractors to bleed the mesh if the rock surface cannot be seen. Traditionally most of the rehabilitation is performed by installing more bolts (Sarpong, 2013).
2.10 Challenges One of the most critical bottlenecks that the Mufulira mine currently faces is that the development is not fulfilling the targets. This affects the entire production and makes it problematic to maintain an even production level. This has in turn resulted in that the mining sequence is not maintained; the top production level should be 3 fans (approximately 5.4m) ahead of the bottom production level in order to get better stress conditions, but since the sequence is not kept the Mine has experienced stress problems in the footwall drifts (Sarpong, 2013).
The main reason that the development is not keeping up with the production is that the mine also struggles with low loader availability, down to 50%. This low availability can further be deducted to
22 problems in getting the right spare parts in time, making the downtime of broken equipment higher than necessary.
Another major bottleneck is the waste handling, as of today all the waste has to be hauled to the surface. This is a major logistical problem which will only grow bigger with the depth (Kampeni, 2013).
Ventilation is also an issue as the mine reaches greater depths the geothermal gradient as well as auto compression will make the air warmer. The mine has also difficulties in getting the fresh air to reach the stopes in the deeper parts of the mine. The main reason for this problem is that in order to recover more ore, old stopes that were sealed have been reopened and mined out; this has led to a major air leakage in the mine. The open stopes also acts as giant radiators and heats up the air. The stopes can furthermore not be sealed since the production is ahead of the development and there is therefore no waste to seal them with. At the same time the fresh air will need to be transported longer distances increasing the risk of air leakages on the way and increasing the capacity demands on the ventilation system (Mutawa, 2013).
On some of the mining levels the mine also experiences issues with water inflow, to tackle this problem some parts of the mine has to be mined using the MCR1 mining method which is much less efficient than the MCR2 mining method as previously described in chapter 1. The water problem also means that in certain parts of the mine pillars has to be left behind so that pumping equipment can be left in these areas. This has led to stress problems in these areas which also have resulted in failures in some of the nearby drifts. Because of this the footwall drifts in the deepest areas have been forced to be moved up to 40-50 m away from the stope (Sarpong, 2013).
2.11 The Future As mining operations continues in Mufulira the demands on the mining operations will increase with the increasing depth. Planning will have to be performed in even closer collaboration with the geotechnical department. The logistics will involve longer transportation distances and will therefore be more costly and equipment will be worn out at a higher rate. The temperature in the mine will rise with the depth due to heating of the air by auto compression and the geothermal gradient. In order for the Mine to maintain an efficient production in the future the Mine has to overcome these issues but also solve its current bottlenecks.
23
3 General description of Mindola mine Mine operations at Mindola mine started in 1935. Mindola shaft is one of the four active mines at the Nkana mine site, located in Kitwe the heart of the Copperbelt of Zambia points out at the map inFigure 12. The three other mines are Central, South Ore Body and North Shafts, all located on the eastern side of a complex folded synclinal structure known as the Nkana Syncline. A sub- vertical shaft has been deepened to expand mining down to a level of 5150 feet (1570m). Currently a new shaft is being planned to take the mine even deeper and extend its life. Figure 13illustrates the location of all four ore sources at the Nkana mine site (Chanda, 2000).
Figure 12 Location of Kitwe, Nkana (Google maps, 2013)
24
Figure 13 Nkana mine site and location of the four ore sources (Bwalya, 2013)
Mindola is owned by Mopani Copper Mines PLC which is a Zambian company owned by Carlisa Investments Cooperation (73,1%), First Quantum Minerals Ltd (16,9%) and ZCCM-IH (10,0%). Mopani directly employs 8000 employees and complements the work force with more than 7800 contractors making the total labor force to be more than 16000 people (in May 2012) (Chanda, 2000).
The total measured and indicated reserve in Mindola is 28 million tones. With an annual production at 1.1 million tons per year the life of mine will be about 25 years. The cutoff grade is at 1% and sub economical grades are about 0.5-1.0%. With sub economical grades it is referred to areas were the copper grade does not meet the cutoff grade nor cannot be classified as waste since it can be beneficial to excavate in order to, for example, maintain an even production (Chanda, 2000).
The minerals extracted at Nkana mining site and Mindola sub vertical shaft are copper and cobalt ore. The mining method used for production at Mindola Shaft is known as the Vertical Crater Retreat (VCR) method. The vertical crater retreat method was first introduced at Mindola Shaft in the late 1980s after several attempts that helped define specific criteria’s for its application. Since then, the method has been improved and updated mainly because of the problems that have arisen over the years. The VCR stope layout is standard in all mining areas with a stope span of 23 m and 7 m rib pillar width. The ore body is on average 12 m wide with dip varying from 56 to 72 degrees and overturned on south. Blasting and drawing is designed so that shrinkage can be used in order to control hanging wall dilution and rib pillar failures. The last ten meters of the stope acts like a temporary crown pillar and is normally broken using choke blasting. The loaders used to draw the ore from the stope are Toro 301 LHD loaders. Production is taking place as far as 2.4 km from the sub vertical shaft (Chanda, 2000).
25
3.1 Geology The ore body at Mindola Shaft forms part of the complex, tightly folded syncline structure, The Nkana Syncline. The syncline comprises east and west limbs, connected by a highly folded complex at the base. It closes towards the south of SOB Shaft and opens towards the north. The syncline generally plunges towards the North-west. The orebodies, average 8 m in width where no folding occurs and dipping to the west at an average of 60 degrees. The Mindola orebody is regular with no folding averaging 12 m in width. Mindola Mine is separated from Central Shaft by the 1.5 km wide Kitwe Barren Gap (Chanda, 2000).
Towards the south from the central position of the mine, the dip becomes steep and over-turned. Towards the north the orebody tends to be shallower and dips about 56 degrees. The orebodies are primarily sulphides. The dominant ore minerals are chalcopyrite, bornite and carrollite. The deposits are contained in a group of sedimentary strata in the lower Roan group of the Katanga System. The Lower Roan is made up dominantly of a footwall formation consisting of archaceous materials, conglomerates, quartzites and sandstone followed by the ore formation of argillite, shales, dolomitic shale and quartzites. Above this are more dolomitic shales with beds of sandstone and quartzite. The orebody is illustrated in Figure 14 (Chanda, 2000).
Figure 14 Mindola ore body (Chanda, 2000)
26
3.1.1 Hydrogeology There are four main aquifers at Mindola Mine; a footwall aquifer and three hanging wall aquifers, these are called Near Water, Far Water and Ultra Far Water. The Footwall water is related to zones of fissuring/joint and leaching, and tends to drain fairly quickly. With caving method of mining, the aquifers, especially the Near Water and Far Water Aquifers, need to be drained below a given horizon before mining can commence. The dewatering should be done five years ahead of production. The dewatering is done by drilling into the aquifers and the water flow is regulated with the diameter of the holes. Pressure gages are installed in the holes in order to know the water level. The near Water is located 50-60 m from the ore body and the Far Water is located 200-250 m away. The Ultra Far Water sediments which lie about 300 m to 350 m above the orebody, and hence, has only been probed bit are not routinely dewatered as it is too far in to the hanging wall. Studies have shown that the situation will remain the same as mining goes deeper (Chanda, 2000).
3.1.2 Exploration
3.1.2.1 Diamond drilling The primary exploration method used in the mine is diamond drilling. It is used to perform the following tasks; (Mopani, 2013d)
Provide advance Geological information on ore distribution and structural nature Using the same geological data to generate new resource base which will be used to replenish future Nkana mine site reserve base. Investigate extensions to the existing ore bodies Explore further ore bodies within the immediate vicinity of the current mine workings Dewatering of aquifers to ensure a flood free mine environment prior to production Special drilling for adverse ground conditions, pilot cover holes when need arises
The diamond drilling is carried out by a contractor and supervised by the mines geologists. The geologists provide the contractor with data regarding drill sites, core size, machine type and other drilling details. When drilling the contractor is expected to follow these standards; (Mopani, 2013d)
Supply core of a minimum of 95 % core recovery. Lower performance entails re-drilling at contractor’ cost. Well stacked, clean and labeled core is delivered to the client’s core shed and signed off. Boreholes in excess of 100m are camera surveyed.
Underground diamond drilling is done from excavations called cubbies which are developed in haulages, water drives, footwall drives or the crosscuts depending on where drilling is needed. In order to generate the ore body an initial drilling pattern is set with a spacing of 150 – 200m between the drill holes. The block generated by the drilling has a vertical height of 50 - 80m. The
27 drilling is centered at 60m in the later stages of mining in order to improve the confidence of the mineral resources. The last step in classifying the mineral resource as an reserve is to drill closely spaced horizontal holes with 30m spacing. All drill cores are collected and logged for rock type, structure, ore minerals and alteration characteristics. The cores are split with a diamond saw and one part is taken to the laboratory for analysis (Mopani, 2013d).
3.1.2.2 Channel Sampling A secondary way of defining the ore resource is chip sampling. The chips are collected inside the orebody and are generally very accurate.
Channel sampling is performed at a spacing interval of 30m in order to correspond to the stope dimensions. Most often only the first crosscut is sampled and chisels are used to collect the samples from the exposed ore in the crosscut. The intervals in which samples are collected is determined by observing the lithology and density of the ore mineralization and the minimum interval is 0.25m (Mopani, 2013d).
3.1.2.3 Sample Handling The validity and integrity of collected samples are very important as a diluted sample will give false readings. Therefore all samples are wrapped in a plastic bag and marked with a tag to make sure the samples are not mixed up before they are delivered to the laboratory. (Mopani, 2013d)
3.1.2.4 Preparation, Analysis and Security A laboratory on site handles copper and cobalt assays and a quality control program is used to ensure the accuracy of the laboratory work. Both ore and milled samples are tested (Mopani, 2013d).
3.1.2.5 Processing and Validation Personnel from the mine are performing the mineral resource estimation for both the underground and open pit operations. Data from both diamond drilling cores and channel sampling is used during the estimation process and the estimation is based on manual polygonal, triangulation and contouring methods. The value for copper and cobalt were determined using weighted grades/tonnage and estimated with planimetered volumes and then converted to a tonnage by using a gravity of 2.56. The fringes of the ore body were determined using a combination of triangulation, grade and thickness contouring (Mopani, 2013d).
3.1.2.6 Grade Control The mine planning department is basing their monthly schedules on data submitted by the geology department. The grades are controlled onsite by personnel from geology.
Monthly estimates of stope tonnage and grades are based on data from working faces and are used to create the schedules. To further control the grades production sampling is carried out and samples are analyzed within 48 hours. Sampling is carried in the following areas;
28
Underground Stope sampling
Conveyor belt sampling
Tray car sampling
Underground Stope Sampling is the most extensive, but also the most accurate way of grade control sampling (Mopani, 2013d).
3.1.2.7 Grade and Tonnage Reconciliation The ore reserves and resources are constantly review and updated by the mine staff. Production and scheduling is evaluated taking into account tonnages, grades and new findings from exploration. Tonnages and grades are reconciled using the statistical method of tracking tonnages and grades back to the source shafts and consequently to stopes and development ends (Mopani, 2013d).
3.2 Geotechnical conditions
3.2.1 Rock mass data for mine design From the lower levels of the Mindola Shaft Rock Mass Classification Data was collected and it showed that the mining rock mass rating (MRMR) of the ore varies between 59% and 79%. The MRMR results represents fair to good ground conditions. The Uniaxial Compressive Strength (UCS) of the orebody generally varies from 40 to 200 MPa. Schistose and low-grade argillites are the weaker rock formations and are amenable to failures at low stress levels. Measurements of the in situ stress conditions have shown that the vertical stress is generally equal to the weight of the overburden and the horizontal stress is about 0.9 times the vertical stress (Chanda, 2000).
At Mindola lower levels scan line joint surveys have been performed and then analyzed using DIPS. The analysis showed three major joint sets in the sampled area. None of the joint sets is the main plane of weakness; it is formed by the bedding in the rock mass. Table 4shows the joint set data. Table 5presents the strength parameters for the different minerals in the Mindola orebody and Figure 15illustrates their relative location (Chanda, 2000).
Table 4 Joint set data (Chanda, 2000) Joint plane Dip Azimuth Comment 1 66 213 Bedding 2 63 48 Joint set 3 72 2 Joint set 4 60 318 Joint set
29
Table 5 Strength parameters for the different minerals in the Mindola orebody (Chanda, 2000) Rock type Compressive Rock Quality Youngs Modules Poission's Ratio Strength [MPa] Designation [ROD] [GPa] Lower Conglomerate 170 80 - 100 76 0.25 Footwall Quartzite 220 80 - 100 66 0.24 Footwall Sandstone 190 75 - 100 70 0.26 Footwall Conglomerate 140 90 - 100 63 0.19 Schistose Ore 50 30 - 40 47 0.38 Low grade Argillite 110 15 - 70 61 0.25 Banded Ore 105 25 - 100 53 0.20 Cherty Ore 170 30 - 100 49 0.18 No.1 Marker 110 40 48 0.25 Porous Sandstone 130 30 - 50 48 0.25 Hangingwall Argillite 190 0 - 50 63 0.21
Figure 15 Minerals in the Mindola orebody (Chanda, 2000)
3.2.2 Ground behavior Ground conditions throughout the lower levels are consistent; in general they do not vary much at all. The factors behind differences in excavation performance are changes in joint distribution and rock strength, the most important factor is the stress loading conditions generated by the mining sequence. During ore excavation the crown pillars, rib pillars and the hanging wall acts as the main
30 supporting structures for the stope. The final phase of stope blasting includes recovering the crown and rib pillars. The ground condition in the crown pillar deteriorates because of increasing stress when stope blasting advances upwards. The increasing stress causes ground movement which further complicates drilling and blasting of the crown pillar. If blasting of the crown pillar is delayed stresses will be induced in the footwall which will affect the stability of the main haulage located in the vicinity of the pillar. In order to secure the stability of the main haulage heavy passive support has to be installed in addition to the standard grouted rebar. In addition to the induced stress in the footwall, seismic events in form of bumping occur when blasting of the crown pillar is delayed. These problems have been solved today when the mining sequence changed to up dip from down dip. This means that there is no need for a crown pillar, eliminating all the problems related to it (Chanda, 2000).
The rib pillars which are located in between the stopes should be recovered by drilling and blasting from the footwall drives. Observations have shown that the rib pillars fracture because of the high stresses. They have also shown that the rib pillars fracture in the direction of the major principle stress, perpendicular to the ore body. The fracturing causes the entire stope to collapse. Recoveries of the rib pillars are very bad or even none existent as the pillars are in yielding conditions. This also affects the overall ore recovery (Chanda, 2000).
The mining excavations can become unstable because of the ground conditions in the abutment area which will result in unsafe working environments especially where stope development is taking place close to the producing stope. In order to prevent ground movement, grouted rebars are installed right after the excavation. It has been suggested by consultants that the high stresses are a result of low degree of caving behind the lower abutment because of the steep dip and narrow orebody. Because of the low caving little or no loads is transmitted to the footwall by the caved material for at least 300m and 600m up dip from the lower abutment. Excavations mined close to the orebody often suffer from severe stress damage. The area within the orebody consists of weak structures which becomes problematic for the VCR drilling chambers when installment of support is delayed. A result from that is caving in the roof, up to 5 m high followed by expensive remedial work in form of passive support (Chanda, 2000).
The induced stress from stope extraction does not only affect sublevel stope development but also main level development. As an example; during main level development the haulage is driven 150 meters ahead of the active stope; this face is in relatively good ground conditions as it only encounters virgin rock stress. However, experience have shown that this stability is only short term because when it advances below the position of the upper level it encounters unstable ground caused by induces stress from the active stope (Chanda, 2000)
In the vicinity of the haulage drifts it is common that collapses occur because of destressing of previously highly stressed rock. It is especially common where grouted rebar support was not
31 installed in addition to steel arches or timber settings during the development stage (Chanda, 2000).
3.3 Mining method The mining method used for production at Mindola Shaft is known as the Vertical Crater Retreat (VCR) method and it is a variation of sublevel stoping. The mining is carried out as shown in Figure 16. Drilling, charging and blasting is carried out from the sublevel drift and the ore is then blasted in slices down into the undercut. Loading of the broken ore takes place in the haulage drift beneath the sublevel drift. The VCR stope layout is standard in all mining areas with a stope span of 23 m and 7 m rib pillar width. The ore body is on average 12 m wide with dip varying from 56 to 72 degrees and overturned on south. Blasting and drawing is designed so that shrinkage can be used in order to control hanging wall dilution and rib pillar failures. The last ten meters of the stope acts like a temporary crown pillar and is normally broken using choke blasting (Chanda, 2000).
Figure 16 VCR stope layout. (Hartman & Mutmansky, 2002)
32
The mining sequence have been changed from down dip to up dip and introduced backfilling because of ground control problems. This change was made to eliminate the stability problems that occurred because of the crown pillars that were left behind when mining down dip, there are no crown pillars left behind when mining up dip (Chanda, 2000).
3.3.1 Main level development From the main levels the ore body is accessed through 6 m wide x 4 m high shaft cross-cuts. Extraction haulages are excavated towards the north and south in parallel with the ore body strike and the drifts are 4 m wide x 3.5 m high. Where mechanized draw point loading is performed the extraction haulages are positioned within a competent Basal Sandstone formation which is located approximately 60 m from the orebody. At this distance the extraction haulages will not be influenced by the mined out stopes and support is sufficient with normal re-bar and cable bolts (Chanda, 2000).
3.3.2 VCR sublevel development The stope development for VCR mining consists of two primary levels, the top drilling level and the draw point level. Where the stope drilling heights exceed 50 m an intermediate level is also established in between the drilling and the draw level. All three levels are accessed by drives from the ramp located in the footwall. Figure 17and Figure 18shows the typical VCR stope profile at Mindola (Chanda, 2000).
Figure 17 Illustration of a typical VCR stope (Chanda, 2000)
33
Figure 18 Illustration of a typical VCR stope (Chanda, 2000)
Access cross cuts are mined at every forth stope to the access drive. The position of the access drive depends on ground conditions and is placed either in the footwall or the orebody. From the drive, drilling chambers are then mined to the hanging wall/footwall. Figure 19 shows the typical top drilling level and rigging positions in the cross-cuts (Chanda, 2000).
34
Figure 19 Top drilling level and rigging positions in the cross-cuts (Chanda, 2000)
Development going through waste formations is 2.4 m wide and 2.4 m wide, just big enough for machines to pass (Raise borer, DTH and Loader). When development reaches the ore body, the height of the roof of the cross cut is increased to 3.5 m. Two different layouts are used for VCR, in the first one a single VCR drilling chamber with a width of 8.0 m and two 2.4 x 3.5 m cross cuts. In the second one two smaller VCR drilling chambers are excavated to about 6.0 m width which results in improved stability for the chamber (Chanda, 2000).
Drilling and blasting of pillars is done from the safety of the footwall drive. For each stope, three 3.5 m wide x 3.5 m high draw point cross cuts are driven from the footwall loader drive. The loader drive is placed 50 m from the ore body in the competent Basal Sandstone below the weak Lower Conglomerate. From the ore body footwall drift, loader cross cuts are then driven through the ore body to the hangingwall. Within the orebody, all the three cross-cuts are shaped to form a single holing chamber (undercut) (Chanda, 2000).
The main considerations in optimizing the layouts are:
stability of the pillars between excavations; cross-cuts are more stable than drives; within the orebody, the hanging wall strata is stronger than the footwall strata; and blast holes should be parallel as far as possible especially near the hanging wall.
At Mindola two different types of drilling layouts are used. In the first one the entire VCR stope is drilled at once and in the second one there is an intermediate drilling level. The advantages with having an intermediate level are increased drilling accuracy, the ability to install support in the hanging wall and to have drawpoints at approximately mid-stope height to improve the draw in the stope. The negative effects of the intermediate level are increased development and stope
35 preparation time. At some levels delays between the preparation of the top drilling level and the intermediate level have led to stopes being pulled down too far below the intermediate level while the tops is still being prepared. This results in damages to the hanging wall beam which will cause and increase in dilution (Chanda, 2000).
In order to keep the development faces clean two different LHD machines are used; the LHD TORO 150D with a bucket capacity of 1.6 m3 and the LHD TORO 301 with a 3.6 m3 bucket capacity. The availability of the smaller loaders is low which has resulted in delays in development of the top levels. Unlike production loaders, development loaders are not hired in terms of maintenance and are very captive. LHD TORO 301 is used for cleaning the intermediate and draw point development faces. During the early stages of VCR development cleaning was done by conventional hand lashing. Before the 150D and 301 LHDs was introduced Micro Scoop loaders where used (Chanda, 2000).
3.3.3 Charging and blasting Previously high bulk slurry was used for sublevel big holes open stoping and VCR blasting. Today 110 mm x 560 mm, 6.25 kg Energex cartridge slurry explosive is used, it very resistant to water and has a density of 1.3g/cc. In order to ensure filling of the hole and to avoid decoupling, the cartridges are slit so that they can easier expand in the hole. The explosive power is concentrated to a length of six times the hole diameter to form a point charge. The technique used for blasting the stopes is known as the cratering technique and each blast breaks about 2.5 m to 3m of the stope back. The holes are plugged using 110 mm diameter x 450 mm length wooden split plugs. For stemming dry and uniform sand is used. In order to use the explosives as efficient as possible the plugging of the holes has to be correctly done as along with the use of both bottom and top stemming and careful positioning of point charges. 25 kg of explosives fill approximately 0.9 m length of the 165 mm diameter blast hole giving a powder factor of 0.3 kg/t of ore. The initiation system used is heavy duty U500 HD Nonel detonators with different tube lengths (21 mm, 30 mm, 45 mm and 60 mm). Each VCR blast gives an average of 2.5 m break and gives about 2 Kt of ore. In most stopes fragmentation of ore is very good giving an average of 0.5 m x 0.4 m x 0.3 m. The charge setup is illustrated in Figure 20 (Chanda, 2000).
36
Figure 20 VCR charge setup (Chanda, 2000)
It is important for the efficiency of the blast pattern that the effects of the crater and cylindrical charges are combined in the outer holes. To be able to make that happen it is of outmost importance that the VCR zone is kept one advance ahead of the stope. After each blast the length of the holes are measured to identify where the ground level in the stope is and to decide which holes should be blasted next. Computer software is used for illustrating the orientation of the blast holes at different depths within the stope. A few specific problems have been encountered during stope blasting; hole closure because of poor ground conditions in some areas, sand baking because of too much stemming and wetness, and blow outs because of insufficient stemming or poor placement of the charge. Another common problem is coning of the hole which makes it difficult to plug the bottom of the hole. If a hole is blocked it is either pumped clean or re-drilled in order to open it up (Chanda, 2000).
Stress has also caused problems by crushing the drill holes, especially in the lower levels and the holes also gets a dog-ear shape up to 3 meters below the floor, probably because of end effect. Because of the crushing the holes have to be cleaned before charging. The stress can also cause hole closure during blasting which can delay the blast or lead to a complete re-drilling of the stope. The final stage of stope blasting is the most critical one, when the final over cut sill is blasted. The length of the sill is usually 8 to 10 m and blasting has been successful in most stopes. The VCR blasting is illustrated in Figure 21 (Chanda, 2000).
37
Figure 21 VCR blasting (Chanda, 2000) 3.4 Mine Planning The Software’s used for planning in Mindola are the package from GemcomSoftwares™ and Microsoft™ Office. Gemcoms package includes SURPAC for design purposes and Mine Sched for midterm planning, long term planning and ore reserves generation. The use of Microsoft Office involves Excel for the short term planning, shift scheduling etc.
3.4.1 Production planning: The mine planning department is planning the production and developing the layout for the mine, based on data from the geology and survey departments. The work procedure for planning the mine is as follows; it starts with a geological section that’s been divided into mining blocks with regard to geological reserves. The blocks are then divided into stopes. The next step is to include the main and sublevel development that is needed to access and extract the stopes. Geological and survey data determines the grades and tonnages for the stopes. The data is also the foundation for the design of VCR development and stope layouts. Planning the stopes also includes developing the drill layouts and including data such as; drilling positions, direction, depth and hole size.
38
Further it also includes calculation of ring and toe burdens and amount of explosives per hole (Chanda, 2000).
Other tasks performed by the planning department are preparation of end of mine life, five-year and one-year production schedules (Chanda, 2000).
3.4.2 Planning challenges The aquifers cause problems as they have to be dewatered before the area can be taken into production. In some areas the ore body is flattening out which causes stability problems as the stope angle should be kept at a minimum of 45 degrees to maintain stability. Before the mining method was changed to up dip, stress concentrations in the crown pillar sometimes induced failures at the footwall side and in the drilling chambers which delayed the drilling and by doing so, it also affected the entire production chain (Chanda, 2000).
3.5 Mine layout and infrastructure The mine is made up of one major ore body that has resources to a depth of 1,800 m. The access to the mine is made through two vertical shafts at ground level; one for ore and one for personnel and machinery. The two upper shafts go down to 880 m and then through a sub-vertical shaft down to 1,600 m. The layout is illustrated in Figure 22and Figure 23 and the drift sizes are presented in Table 6 (Chanda, 2000).
Figure 22 Layout of Mindola underground mine (Bwalya, 2013)
39
Figure 23 Layout of the new planned levels (Bwalya, 2013)
40
Table 6 Drift sizes (Chanda, 2000) Drift size Footwall Access Drive - 4.0mW x 4.0mH Top Access Cross Cuts - 3.6mW x 3.5mH VCR Hangingwall Drive - 4.0mW x 4mH VCR Drilling Chamber - 4.0mH-width varies Loader Drive - 4.0mW x 4.0mH Ore Pass -2.4m diameter
3.5.1 Hoisting system After the ore is blasted it is draw from the stope using the TORO 301 load-haul-dump (LHD) machines. The draw point consists of a loader drive, two or three draw point cross-cuts (depending on size of stope), a tip and a ventilation return airway. Each production stope in the retreat has one LHD assigned to it. The target for each LHD is to average about 40 Tonnes per hour and most of the production takes place during the afternoon and night shift. Low availability on development loaders and DTH machines is at times causing delays in stope preparation which in addition is reducing the productivity of the production loaders. Another factor that impacts on the productivity is long tipping distances from the draw point to the haulage drift. The distances that the loaders have to travel are between 50 m and 150 m. Availability of the hired loaders is high, the only constraint is availability of ground at times, which results into low utilization of the machines (Chanda, 2000).
The ore pulled from the stope is tipped down into a 24 m long, 2.4 m diameter raise bored ore pass. In the bottom of the shaft a 2.4 m box is installed for filling the trays. The boxes are operated through the use of highly pressurized air. The ore is then loaded into trays carrying 12 tons of ore each. The trays are pulled by 15 ton Jeffery and Batman Electric Locos. The locos transports the ore to the crusher it is crushed and then loaded into the skips using a conveyor system. The ore is skipped to the old crushing level where it is further transported, by the use of conveyors, to the next skip which takes the ore to the surface for further processing. The hosting system is illustrated below in Figure 24 (Chanda, 2000).
41
Figure 24 Hoisting system (Bwalya, 2013)
3.6 Ventilation The ventilation system in the Mindola mine is based on the technique that creates a pressure difference between the mine and the surface. This is done by having up cast fans that sucks out the air from the mine and by doing so creating a lower pressure within the mine which makes the fresh air seek its own way down in to the mine through the main access shafts, a simplified ventilation flow chart are presented in Figure 25 (Mutawa, 2013).
42
Figure 25 Ventilation system in Mindola. Three up-cast fans illustrated on top. 3.7 Geotechnical considerations
3.7.1 Failure The most common failure mechanisms are slow and time dependent deformations and in some cases stress concentrations in combination with unfavorable rock structures. Another known mechanism is stressing and destressing of areas in the footwall which results in fallouts. The magnitude of the failures is most often small, local and stress induced e.g. crushing in the vicinity of the crown pillars. It is very rare that the failures delay the production in any larger extent (Katongo, 2013).
3.7.2 Seismicity The seismic events in the mine are triggered by the mining operation and primarily consist of sliding in the hanging wall and failures related to the stress concentrations in areas close to the crown pillars. The issue with stress concentrations related to the crown pillar will be eliminated once the old down dip mining areas are mined out (Katongo, 2013). In order to monitor the seismic events and activities in the mine a seismic monitoring system will be installed. The idea behind the system is to gather data that will allow the mine to predict where and when a seismic event will occur (Katongo, 2013). It is not common that seismic events occur but when they do the mine has a strict procedure to follow: (Mopani, 2013e)
1. When the event is heard/felt, all persons in the area should withdraw to a safe places and remain there, and should not run about, until the seismicity dies out.
43
2. The most senior employee on site shall withdraw his men in a quick but orderly manner from the working place. A safe area shall be any place sufficiently far enough from the mine workings not to be affected by seismic event (preferably in excavation within FW Sandstone/Quartzite formations). The Supervisor to ensure that the safe area to be free from bad hangings and to have adequate ventilation.
3. Upon reaching a safe area, the most Senior employee shall carryout a census of men to ensure that no one is missing. In case the route is completely covered by fall of ground from the seismic event making escape to a safe area outside the working place impossible, the Supervisor shall move his men to a safe place within the closed-off area and wait until the route has been made passable.
4. If no one is missing the senior most employee shall lead his men to the nearest waiting place. No personnel shall be allowed to enter the area affected by a seismic event for at least 8 hours after the seismic event occurrence, except for rescue purposes.
5. If it is discovered that some workers are missing, the most senior employee shall lead his men to the nearest waiting place while he will rush to the nearest telephone to inform the Shaft Control Office and his Mine Captain about the missing men.
6. The Shaft Control Office shall then inform the senior officials, the UGM, the Senior Safety Officer, the Senior Geotechnical Engineer. An inspection of the area affected by the seismic event shall be carried out within 24 hours by a team comprising at least of a Geotechnical Engineer, a Safety Officer and the Mine Captain for the Section. If the area is still seismically active 8 hours after the seismic event occurrence, the Geotechnical Engineer/underground manager shall decide to postpone the visit until the seismic activity has died down.
7. In the event of any person(s) being trapped by the fall of ground, the UGM shall put in place a team and give instructions to start the rescue operations.
8. Upon inspection, the Geotechnical Engineer shall give instruction to the Mine Captain whether or not to commence mining activities in the area depending upon the observed Geotechnical conditions (including support work). The instruction will be given in consultation with the Safety Officer and the Mine Captain.
9. Such instructions shall be confirmed to the Underground Manager in writing before operations recommence in the area. Rehabilitation and support recommendations shall accompany the instructions to recommence mining activities.
44
10. The Geotechnical Engineer shall write into the area statutory log book the instructions given underground on whether to recommence mining activities or not and the rehabilitation and support recommendations given.
3.8 Ground control practice
3.8.1 Damage mapping There is no established system or database for damage mapping, it is performed by the geomechanic engineers when they are down in the mine performing inspections on rock support and after an event has occurred. It is important to determine the cause of the damage and which failure mechanism is causing it. When indentifying the failure mechanism observations, experience, calculations and numerical models are taken into account (Katongo, 2013).
3.8.2 Numerical analysis The software’s used for creating numerical models are Phase2, Unwedge and Examine2D. The numerical models are used for various purposes:
Mine-wide analyses are made with Examine2D and it was used to design the up dip mining sequence and its echelon form For larger constructions like workshops, structures in the rock mass are analyzed with Unwedge. In case of a failure models can be used to identify the failure mode and what caused the failure. Factors considered when making the model; Location of current mining, Was there any blasting nearby etc. For design of rock support models are used initially to see the extent of stress zones in order know how long the bolts have to be and which stress situation that can be expected. All the support is standardized so no comprehensive analysis is needed at this moment.
In order to make the numerical models as accurate as possible rock parameters like RQD, RMR, GSI and joint mapping are used as input in the models. In some area squeezing conditions have to be taken into consideration (Katongo, 2013).
3.8.3 Support standards The rock support standards are developed by the rock mechanics crew and are based on experience and numerical analysis with careful consideration of the mechanical properties of the rock mass and its stress situation. The support standards are presented in Appendix B (Katongo, 2013).
The most common support is rock bolts, consisting of rebar, fully grouted with ordinary cement. When development has to be rushed hydrabolt is used (friction bolt similar to Swellex) or the ordinary cement is replaced with fast setting cement. Tendon straps are used where there are high
45 stresses and where the standards require it. Shotcrete is used in access drifts and inside the stope during the final stages of production. It is also used in some areas as a psychological factor. Diamond Wire mesh is also used as surface support (Katongo, 2013). The support requirements for different excavations are divided into three categories: (Katongo, 2013) 1. Big sized constructions like workshops and pumping facilities that are placed far away from the production area. Support: Cable bolts and if needed shotcrete. Rock mass: Fair to Competent rock mass 2. Primary development like the ramp and tramming drifts. Support: Roof bolts and mesh if needed Rock mass: Fair to Competent rock mass 3. Stope and footwall development Support: Roof bolts Before drilling is commenced initially temporary support is installed at the front after scaling and cleaning of the rock. The temporary support consists of a fine mesh and timber piles. Further support is standardized with different classes depending on which kind of drift that are being supported. The following is a description of the development support standards: (Mopani, 2013f)
Manual scaling at start and during shift
Mechanical props to be used during Jack Hammer support hole/face drilling and charging
Do not enter unsupported areas
At least the next stope development should be fully supported prior to commencement of blasting the preceding stope.
Should ground condition adversely change or emergency of weak Discontinuity planes, Rock Mechanics personnel must immediately be informed.
Ground conditions around the VCR Drilling Chamber may deteriorate due to stress loading and blasting, as blasting progresses up-dip, hence passive support in form of 04 pre- stressed timber packs should be installed before blasting commences.
Roof bolt: 2.4m long 12-14 ton capacity, 500mm minimum thread length, c/w 150mmx150mmx4mm thick, dog-eared face plate, installed on square grid
Maximum allowable bolt protrusion is 50 -100mm, for both roof bolt and threaded cable bolts, and 500mm for plain cable bolts. All shallow dip holes should be flashed out of chippings before grouting roof bolt
46
Section officials are directly responsible for enforcement of support standards (compliance & quality)
No rock should fall uncontrolled
After the support is installed it is visually inspected and in some cases pull out tests are performed. What is being looked at during the inspection is if the support has been installed according to the standard used in the area. As an example: If the roof bolts are not installed perpendicular to the rock surface the contractor won’t get paid for the work and will have to do it again according to the standard. The support standards are presented in Appendix B (Katongo, 2013).
3.8.4 Rehabilitation No system is implemented for decision making regarding rehabilitation of support. Rehabilitation is performed when the rock mechanics engineers consider it necessary. Which kind of rehabilitation that is needed is based on experience and judgment by the rock mechanic engineers. It also factored in which area that is in need of rehabilitation (Katongo, 2013).
3.9 Challenges In order to fulfill the future goals of the mine and continue production at greater depths a new shaft is planned and construction will start in a near future, development should be done in 2016. The shaft will be 1.9 km deep and will bring the mine down to 5,986 ft where a new main level is to be constructed. No problems related to stresses are expected as the rock is very competent and the geology is very favorable as it is consistent down to great depths. There will be two crushers at this level and ore can be hoisted directly to the surface without being handled at an intermediate level. This new level will be used as a main level far into the future. If mining proceeds even deeper it is being considered to use conveyors to hoist the ore to the shaft. When the mine deepens the workload on the ventilation system increases. The air have to travel further and the impact of the thermal gradient and pressure heats the air so at a certain point the air temperature will be so high that refrigeration is needed. Today the ventilation system is sensitive; each of the up cast fans is connected to a shaft and a specific area of the mine. This results in that when the fan for a curtain area is not functioning, that area will not get any fresh air and production will be interrupted. In the future the down cast shafts will be enlarged so that they can easier supply more air to the mine. New ventilation shafts are planned in order to increase the flexibility of the ventilation and two new raiseborring machines have been ordered to speed up the process. The new shaft that will deepen the mine will also provide an additional large down cast shaft. In order for Mindola mine to maintain a high production at as low cost as possible their current issues and bottlenecks needs attention. There are two critical things that can be done in the near future at Mindola, the first is careful control of the development and production to ensure that
47 they stay at an even level which will make the process flow smoother and more effective. The second is that maintenance of mining equipment has to be improved in order to cut costs added by malfunctioning equipment. This will be ever so important when the mine reaches greater depths where the travel distances will be longer which results in higher wear on the equipment.
48
Part B
Challenges
49
4 Stress The stress conditions within a rock mass before it is disturbed by excavations are related to the weight of the overlaying rock and the geological history of the rock mass. It is then inevitable that when excavations are located at greater depths they encounter high stress environments. These environments create great challenges for the mining companies as their mining operations progress. In order to overcome these challenges the engineers planning these mines have to consider every aspect of the problem when designing the mine. One of the major and most important decisions that they will face is how to extract the ore, which mining method and mining sequence to utilize. 4.1 Numerical Analysis During our visit to Mufulira it was evident that they are experiencing stability problems in their footwall drifts and it is directly related to their mining method. The MCR mining method utilized in the Mufulira mine leaves behind crown pillars which tend to act as stress windows affecting the stress levels in the footwall. As a result of this the mine has moved their footwall drifts further away from the stopes; up to 50m. These problems could be eliminated if the mine would make a decision to change the mining method to one that does not leave behind any crown pillars. Figure 26 gives an illustration of the relative position of the crown pillar and footwall drifts.
50
Figure 26 Illustrations of relative positions of the crown pillar and the footwall drifts.
The numerical analysis presented in this report was based on data collected from Mindola and Mufulira Mines and was performed using the finite element software Phase2. The model was based on a model received from Mufulira where all the inputs where already in place. The input parameters are presented in Table C 1toTable C 4. The first step was to verify that the inputs where correct and that the model itself was valid for our analysis. Two different models were set up in order to model the intended scenarios. The first model represents the first scenario utilizing the down-dip mining method where crown pillars are left behind and the second model represents the second scenario which utilizes the up-dip mining method where no pillars are left behind and the stope is backfilled with waste rock. A third scenario, where the foot wall drift was moved closer to the stope, was also tested using the second model. The excavation stages and the mining sequences used in both models are illustrated in Appendix D. Stress levels were measured in the roof of the footwall drifts for all three scenarios.
51
4.1.1 Model setup The model was set up to represent the mine from 1000 meters depth, past the current mining level at 1540 meters and down to 1850 meters depth which will be an active mining area in the future. The analysis of the model did mainly focus on three drifts below 1700 meters but did also take into consideration the other parts of the mine. The chosen drifts would be located at 1735m, 1803m and 1853m depth. The model is illustrated in Figure 27.
Figure 27 Illustration of the model used in Phase 2D. The area inside the red square represents the analyzed drifts.
The first step in creating the model was to set the external boundaries. The boundaries should be so far from the modeled mining area that they do not affect it in any way. When the outer boundary was in place the next step was to include the orebody and the drifts needed for the analysis. The orebody was then divided into stopes by adding material boundaries. Now the next step was to add the material properties to the different areas and in this model there are four different material properties; hangingwall, footwall, ore shale and backfill. When the material properties where in place, loading conditions were applied to the model and in this case gravitational force was used. The model was discretized1 and meshed2 and the mesh density was
1 The discretization of the model boundaries, forms the framework for the finite element mesh, and is indicated by small red crosses subdividing the boundary line segments. Each cross indicates the position of a finite element node on the boundary.
52 increased around the orebody in order to get better results. After all the inputs where in place the mining sequence was included by excavating the stopes in steps. When the sequence was considered to be a close representation of reality the model was calculated and interpreted. Material queries where added to the roof of the drifts analyzed so that the stress could be monitored as mining progressed within the model.
4.1.2 Crown pillar This analysis and its results are used as a basecase, since it represents the situation in Mufulira today. In Mufulira they are using the MCR mining method which is a method that leaves behind crown pillars. The footwall drifts are placed 50m from the stopes and there is no backfilling of the stopes.
4.1.3 Backfill This analysis was performed with the objective to investigate the effects on the stress levels in the footwall when changing mining method in Mufulira, from MCR to a sublevel stoping method with backfill. In the model, the stope was filled with a rock fill with strength parameters representing waste rock.
4.1.4 Footwall drive distance from stope For this analysis it was assumed that the mining method had been changed from MCR to a sublevel stoping method with backfill where no crown pillars were left behind. The objective of the analysis was to determine if the footwall drive could be moved closer to the stope without experiencing higher stresses than with the current distance of 50m.
4.2 Results
4.2.1 Crown pillar The most critical steps in this analysis was when the stope above the drift was mined out and development on the same level had started, this step gave us the highest stress measured in the drift. Therefore it will serve as the reference value when comparing the results from the other assays. The highest stress value received in the targeted area was126MPa, shown in Figure 28.Figure D 1 to Figure D 9 in Appendix D shows how the stress condition changes as the mining operation progresses.
2 The finite element mesh is created based on either triangular or quadrilateral finite elements. The mesh can, for example, be programmed to contain the material and structural properties which define how the structure will react to certain loading conditions.
53
Crown pillar drift 50m from the stope 140
120
100
80
1735m
60 1803m Stress [MPa] Stress 1853m
40
20
0 0 2 4 6 8 10 Excavation stages
Figure 28 Stress level measurements with drown pillars and the footwall drift located 50m from the stope.
54
4.2.2 Backfill The analysis indicates that changing the mining method would lower the stress affecting the foot wall drifts. When comparing the received values with the basecase values it shows that the highest stress levels in the model have reduced from 126MPa to90MPa.The stress levels in the targeted drifts during the excavation of the stopes are presented in Figure 29. Figure E 1 to Figure E 8in Appendix E shows how the stress conditions changes as the mining operation progresses. Changed mining method Footwall drift 50m from the stope 100
90
80
70
60
50 1735m 1803m Stress [MPa] Stress 40 1853m
30
20
10
0 0 1 2 3 4 5 6 7 8 9 Excavation stages
Figure 29 Stress level measurements with no crown pillar and footwall drift located 50m from the stope
55
4.2.3 Footwall drive distance from stope After running different scenarios and comparing the stress levels to the basecase values, it was clear that the footwall drift could be placed at a distance of 30 meters from the stope and still have approximately the same stress levels as today. The stress levels in the targeted drifts during the excavation of the stopes are presented in Figure 30. Figure F 1to Figure F 8 in Appendix F shows how the stress condition changes as the mining operation progresses.
Changed mining method Footwall drift 30m from the stope 120
100
80
60 1735m
1803m Stress [MPa] Stress 1853m 40
20
0 0 1 2 3 4 5 6 7 8 9 Excavation stages
Figure 30 Stress level measurements with no crown pillar and footwall drift located 30m from the stope
56
4.3 Conclusion After performing the numerical analysis of the identified problems in Mufulira it can be concluded that changing mining method would be beneficial from a geotechnical point of view. The change would completely eliminate the stability issues related to the crown pillars.
The analysis performed also indicates that changing the mining method could have positive effects on the production since it would allow the mine to place the footwall drifts closer to the stope; at a distance of 30m instead of 50m. By placing the footwall drifts closer to the stope the crosscuts would be shorter resulting in reduced time, less recourses needed for development and reduced cycle times for the production. Furthermore, the distance that the equipment travel would be reduced which would result in lesser wear on the equipment and lowered fuel consumption.
The introduction of backfill would remove the need to hoist waste rock to the ground surface as the waste would be used for backfilling the stopes. In the future when the mining operation reaches greater depths the distances to the surface will be even longer and it will prove to have been advantageous to fill the stopes with waste instead of hosting it to the surface. On the other hand the backfill operation could create a need for complementing the machine park in order to be able to move the waste to the stopes. Another issue that could also arise is shortage of filling material if the development is not generating enough waste rock which would result in either increased development or a need to transport waste from the surface to the stopes.
57
5 Ventilation A mine without ventilation would be a very hostile environment for humans; therefore mining ventilation plays a vital role in the everyday mining operations. The role of mine ventilation as a part of mining operations has been dated all the way back to the first millennium BC (Hartman H. L., Mutmansky, Ramani, & Wang, 1997). Historically the first aspect of mine ventilation was to provide fresh air sufficient enough to replace the oxygen consumed by the mine workers (de la Vergne, 2003). Today mining ventilation is considered to be an art of mining and is used to control the mines atmospheric temperature, humidity and dust levels as well as to dilute the noxious gases produced by exhausts from mining equipment and by blasting. Providing enough oxygen for the mine workers has almost become a secondary aspect since the flow of fresh air required to diluting the noxious gases or reducing the atmospheric heat generally is much higher than that what is required to only replace the oxygen for the miners (Pareja, 2000).
5.1 Heat Already in 1949 Spalding (Spalding, 1949) stated that high rock temperature was the factor that most likely would limit the depth of which mining operations could extend to. Hartmann et al. (Hartman H. L., Mutmansky, Ramani, & Wang, 1997) further noted that the demands on the ventilation systems as well as the expenses it would imply to meet them would reach unsustainable levels at great depths.
Heat as an air contaminant is much different from other contaminants such as gases and dust which ventilation system originally was designed to take care of. The main strategy when designing a ventilation system is usually to dilute the contaminant, but heat cannot be diluted. Other measures of ensuring air quality can be to use a respirator to clean the inspired air but nor will a respirator remove heat. Heated air can however be `re-conditioned’ meaning that heat can be removed through an air refrigeration process or by air conditioning (Brake & Fulker, 2000).
The level of which the heat is experienced by humans can also be reduced by increasing the air velocity over the human skin. The wind speed is a key factor for the rate of which sweat can evaporate and thereby cool the human body. This technique is often one of the cheapest solutions to lower the temperature a few degrees but it has a compelling limitation. At a certain depth denoted as the critical depth the air solely due to autocompression no longer has any cooling potential and the only option to lower the temperature that remains is therefore to use air conditioning (Brake & Fulker, 2000).
5.1.1 Autocompression Autocompression can be explained by the way gas in a compressor reacts, when the air enters the mine through a shaft the air is compressed and heated as it is transported downwards since the potential energy of the air is converted in to thermal energy. If no heat is exchanged during the
58 process and the moisture content remains the same; the compression will occur adiabatically following the adiabatic law (Hartman H. L., Mutmansky, Ramani, & Wang, 1997).
5.1.2 Critical depth and critical rock temperature The critical depth is determined only as a function of surface temperature and the depth of the mine, therefore the critical depth varies with the seasons. Above this depth the cheapest solution of lowering the temperature may very well be to flood the workings with air. Doing the same thing below the critical depth gives the exact opposite effect since the human body will gain heat through convection and radiation from the hot air. At this depth the air itself adds to the heat load and the airflow should therefore be kept at a practical minimum. The method of controlling the heat should revert to refrigeration (Brake & Fulker, 2000).
At the same time Hartman et al. (Hartman H. L., Mutmansky, Ramani, & Wang, 1997) denotes that the host rock mass of the mine due to the geothermal gradient almost always is the main source of heat. The geothermal gradient is the gradient of which the rock temperature rises with the depth. The gradient varies from location to location and depends on the thermal properties of the rock and the proximity to igneous activity (Hartman H. L., Mutmansky, Ramani, & Wang, 1997). The geothermal gradient in the copperbelt region of Zambia is approximately 1.8oC/100m depth (Mutawa, 2013).
If rock heat alone were to govern the air temperature the critical rock temperature at which the air would have lost its cooling potential is considered to be around 41 oC (Hartman H. L., Mutmansky, Ramani, & Wang, 1997). Figure 31 illustrates that this temperature often is reached at far less depths than the critical depth of autocompression (Hartman H. L., Mutmansky, Ramani, & Wang, 1997). It must however be noted that in the presented study the assumed surface rock temperature was only 18oC this can be compared to the surface rock temperature of around 25oC (Mutawa, 2013) for the deep African mines presented in this report. The plotted curve for autocompression can therefore be slightly adjusted upwards meaning that the critical rock temperature for these mines will be reached closer to the critical depth.
59
Figure 31 Average geothermal gradients of various worldwide mining districts plotted against the effects of autocompression on o o air wet-bulb temperature as a function of depth below surface. (Assumed surface tw = 18 C and Δtw = 4.4 C/1000m). (Hartman H. L., Mutmansky, Ramani, & Wang, 1997, p. 591)
5.1.3 Heat sources The heat in a deep mine can have many sources, Hartman et al. (Hartman H. L., Mutmansky, Ramani, & Wang, 1997) listed nine potential sources of heat in mines: Autocompression, wall rock, underground water, machinery and lights, human metabolism, oxidation, blasting, rock movement and pipelines. Out of this wall rock, autocompression, underground water and mining equipment are the most capable of creating unacceptable atmospheric heat levels (Hartman H. L., Mutmansky, Ramani, & Wang, 1997).
It needs however to be understood that the level of effect the heat source can have varies greatly from mine to mine depending on characteristics such as; layout, depth, geographic location, equipment selection and mining method among others. The following illustrations underline this statement; Figure 32 illustrates the average heat sources for seven North American underground
60 mines and Figure 33 illustrates the determined heat sources at Creighton Mine, Ontario Canada 1993.
Figure 32 Average heat sources for seven hot North American mines. (Pareja, 2000)
Heat sources at Creighton mine, 1993
Auto compression
11% 19% Wall rock
20% Blasting and 48% metabolism Equipment
2% Underground water
Figure 33 Heat sources at Creighton Mine 1993. (Pareja, 2000)
It should be noted that in Figure 33 broken rock is included in the category wall rock. The broken rock as a source of rock heat can be substantial and can stand for up to 60% of the total rock heat inflow (Hartman H. L., Mutmansky, Ramani, & Wang, 1997). Measures of controlling this heat would of course have to be different depending on the mining method and ore body geometry as well as production layouts, but the principal solution would be to reduce the volume of broken rock and/or increase the hauling rate of it. (Deep underground Hard rock mining) Mathews (1989)(Deep underground Hard rock mining) also points out the heat flow from rock walls as the major heat source, standing for up to 50% of the heat flow in the stoping areas of deep mines. He
61 further denotes that most often this heat exchange is carried out when the air ventilates mined-out areas. In order to reduce this heat flow and also improve the ventilation system he proposes the use of backfill in the mined out stopes (Pareja, 2000).
5.1.4 Physiological heat effects on humans The thermostatic control system in the human body is remarkably exact in holding the body temperature nearly constant at 37.0°C. This system balance heat-loss and heat-gain processes to prevent harmful heat strain on the human body. If the body is overheated heat flows from the warmer mass to the cooler mass; the skin which usually is at least 1°C cooler than the rest of the body. The heat is then discharged through one of the conventional heat transfer processes; convection, radiation and/or evaporation (usually of sweat) (Hartman H. L., Mutmansky, Ramani, & Wang, 1997). The ability of the human body to discharge the heat differs from person to person and depends on the following factors: Cardiovascular fitness, Obesity, Acclimatization, Hydration levels and Clothing (Brake & Fulker, 2000).
The ability to dissipate heat also depends on the environmental conditions; the mechanism that the body employs to cool itself is to sweat but in order for this process to cool the body the sweat needs to be able to evaporate. Naturally the easier it is for sweat to evaporate the more effectively the heat from the human body is discharged. The two most significant factors affecting this process are humidity and air speed over the skin. The dryer the air is the easier it is for sweat to evaporate, that is why hot humid air is more stressful for the human body than hot dry air. At the same time the rate of evaporation increases with the airspeed under all circumstances except when the air has a humidity of 100% or very high wet bulb temperature; in such conditions the effect of evaporation will be less than the effect of radiation and convection and the human body will instead gain heat (Brake & Fulker, 2000).
If the wet bulb temperature exceeds 32°C the human body is subjected to heat strain and work output and safety suffers (Hartman & Mutmansky, 2002). According to the Mine Safety and Health Administration in the U.S. the lowest accident rates are related to working environment with temperatures below 21°C and the highest accident rates to temperatures above 27°C (de la Vergne, 2003). If the working environment cooling power is exceeded by the metabolic rate of the human body the heat cannot be discharged and can quickly lead to dangerous internal temperatures (Brake & Fulker, 2000). This condition is called heat exhaustion if a human begins to suffer from heat exhaustion the early warnings are usually; Headache, Nausea, Dizziness, Weakness, Irritability, Thirst, Heavy sweating and High body temperature. Already at this stage productivity and ability to make decisions are seriously decreased. If the early signs of heat exhaustions are taken seriously it can usually be treated by simple solutions such as resting, resorting to cooler environments and drinking water. If however this condition is ignored the human is in the danger of suffering from heat stroke with much more serious symptoms such as
62
Loss of consciousness, Seizures, Hot and dry skin, Confusions or delirium, this conditions can be life threatening and medical attention should be sought immediately (MSHA, 2013).
5.2 Ventilation techniques As already implied; there are two basic techniques of controlling the temperature in a mine either by flooding the mine with air through mechanical ventilation or by the use of air conditioning or both. Mechanical ventilation will always be needed to dilute noxious gases released in the mine atmosphere during production. If too high airspeeds are required to control the temperature or if the air reaches temperatures in the range of the human body mechanical ventilation should be complemented with air conditioning (Brake & Fulker, 2000).
5.2.1 Mechanical ventilation Mechanical ventilation is responsible for the quantity control of the air in the mine, its movements and its distribution (Hartman & Mutmansky, 2002). When designing mine ventilation system certain factors have to be determined; which activities that require ventilation, the quantity of air required for the various activities, the amount of head losses due to friction in the airways and the actual percentage of fresh air that will reach its destination (Brake & Fulker, 2000). Simplifying there is basically three ways of getting the air down in the mine; either you install downcast fans that push the fresh air down in to the mine or you install up cast fans that creates a lower pressure in the mine resulting in that fresh air will be sucked down in to the mine at all entrances, or lastly you employ a combination of the two methods. The method of installing up cast fans are generally most employed since it is more energy efficient and also since the risk of adding heat to the intake air from the fans are avoided. Directional fans down in the mine can also be installed to help guide the fresh air to its destination.
Having an inefficient ventilation system can be very expensive since the ventilation can stand for 40% of the mines total electrical power consumption (de la Vergne, 2003). To ensure efficiency of the ventilation system air leakages and head losses in the airways should be kept at a minimum. In fact the importance of airways cannot be stressed enough; in general approximately 60% of the all the surface intake and exhausts fans motor capacity is used to just overcome the friction in the airways. Each of those horsepower’s lost to friction is further converted into heat underground. Furthermore if the ventilation shares shaft with the production the resistance is generally five times higher than in a clean ventilation shaft and a ventilation shaft lined with concrete lining has 4 times less friction than a raw raise of the same dimensions (de la Vergne, 2003).
Even though a critical depth exists of which beyond this depth the air movement should be held at a practical minimum; there also exists a lower limit of airflow that must be achieved. (Brake & Fulker, 2000)Illustrated, see Figure 34, the relationship between the thermal work limit (TWL), measured in watts per square m body surface, and the air velocity for higher wet bulb temperature. Studying Figure 34 it can be seen that already at a level of 0.5 to 0.7 m/s most of the benefits of
63 higher airflows is harnessed. (Brake & Fulker, 2000) Hartman et al. (Hartman H. L., Mutmansky, Ramani, & Wang, 1997) also denotes that adequate ventilation should be provided to enhance heat loss, particularly by evaporation but that excessive velocities at the same time should be avoided if air temperature is above the human body temperature due to heat gain from convection and radiation (Hartman H. L., Mutmansky, Ramani, & Wang, 1997).
Figure 34 Thermal work limit vs. air velocity. (Brake & Fulker, 2000)
5.2.2 Air conditioning Due to the relentless and compelling effects of autocompression and regardless of how effective the mine ventilation system is; at a certain depth air conditioning will be required (Hartman H. L., Mutmansky, Ramani, & Wang, 1997). With the term Air conditioning it is referred to the technique of controlling both the air humidity and the air temperature. The most frequently used air conditioning technique is to refrigerate and to dehumidify the air. Dehumidifying is very important since air that is humid puts a greater heat load on the human body than dry air.
A common way of refrigerate is to use the so called change of state or vapor refrigeration (Hartman H. L., Mutmansky, Ramani, & Wang, 1997). In this system the refrigeration liquid absorbs and discharges heat by alternating between liquid phase and vapor phase. Combined with this it is
64 common to use cooling towers, these can be placed either at the surface or underground and is basically a heat-exchange device that cools the condenser water in the refrigeration system. The inflow air is then generally cooled in the way that the air is blown either through a spray of cold water from the refrigeration plant or over cold coils. Figure 35 illustrates a general schematic of a heat transfer circuit (Hartman H. L., Mutmansky, Ramani, & Wang, 1997).
Figure 35 Schematic of heat-transfer circuits utilized in a conventional mine air conditioning system. (Hartman H. L., Mutmansky, Ramani, & Wang, 1997) 5.3 Identified problems Based on the information gathered for the two case studies and the information gathered in the literature study; it is evident that the demands on the ventilation increases with the depth of the mine. The major problem in both mines was to lower and to control the atmospheric temperature in the production areas at the deeper levels. Temperatures in the stoping areas in Mufulira have been measured up to 36°C wet bulb which is much higher than what is recommended from a health perspective. It should be added that none of the mine experienced any problems in diluting noxious gases or keeping dust levels low.
As explained in the previous chapter the heat sources can be many. No official identification of the heat sources has been made in Mindola and Mufulira but the major heat source was considered to be wall rock in both of the mines. In Mufulira which experienced more heat problems than Mindola this could further be narrowed down to the mined out stope walls being the main contributor of the heat inflow. The reason for this is that in Mufulira many of the old mined out stopes have been reopened. The consequence of this is that the fresh air detours in to the old stopes instead of going directly to the destination at the production. The mined out stopes then acts as giant radiators and heats up the air. So at the time the air reaches its destination it has lost parts of its original velocity and is already heated. A simplified scheme of the ventilation problems experienced in Mufulira is presented in Figure 36.
65
Figure 36 Simplified scheme over the ventilation problems in Mufulira.
5.3.1 Solutions for the near future Due to the increasing heat load in the deep parts of the Mine measures to solve this problem has been investigated and undertaken. The measure currently undertaken is to seal the mined out stopes. The ventilation rises between the production levels are also being expanded from 1.8m to 2.4m and the ventilation shafts are going to be expanded from 4m to 6,2m in diameter. The exhaust fans will further be upgraded and extra fans have been installed in the lower levels in an attempt to try and guide the fresh air to the stopes. The new exhaust fans will have an effect of 14kPa each and pump out 650m3 of air per second. This will increase the airflow from 25m/s in the exhaust shaft to 50m/s. With drifts that measure 4x4m this would imply an increase in airflow velocity from approximately 1.6 to 3.1m/s. It is expected that each m/s increase of the airflow velocity corresponds to an approximate lowering of the air temperature with 2°C. This means that the increase in airflow velocity would result in a lowering of the temperature with approximately 3°C.
With the construction of the new shaft, the ventilation system will furthermore be improved since the shaft means that two separate ventilation systems can be installed so that a breakdown in the ventilation won’t affect the entire production. With these improvements the Mine hopes to be
66 able to take care of the ventilation problem and meet the work environment regulations of a 31°C wet bulb temperature.
5.4 Analysis The ventilation challenges that the two mine experience are both related to the massive heat load experienced due to the increasing depth. Both mines identifies the wall rock to be the main source of heat, due to the complexity of heat transfer to the mine atmosphere no calculations to determine the actual load has been made in this report. The analysis has instead been focused on the already undertaken and the proposed solutions from the mines to these problems. The objective with this analysis was to validate if the presented solutions in fact would solve the problems or not.
As discussed in chapter 5.1.2 a critical depth exists of where the air due to autocompression loses its cooling potential and that any air circulation below this depth adds to the heat load instead of lowering it (Hartman H. L., Mutmansky, Ramani, & Wang, 1997). As Hartman et al. (Hartman H. L., Mutmansky, Ramani, & Wang, 1997) also pointed out that the critical rock temperature often is reached before the critical depth also this factor was calculated.
5.4.1 The critical depth due to autocompression In order to calculate at which depth the critical depth due to autocompression would occur a wet ° bulb threshold limit of twl= 26 C (Hartman H. L., Mutmansky, Ramani, & Wang, 1997) was used. If the air reaches higher temperature than that the air adds to the heat load on the human body instead of cooling it. The depth at which this temperature would be reached was then estimated using the following relationship by Hartman et al. (Hartman H. L., Mutmansky, Ramani, & Wang, 1997)
Where; Δtw= wet bulb temperature and ΔZ= the difference in depth.
5.4.2 The depth of the critical rock temperature The depth (z) of which the critical rock temperature of 41°C (Hartman H. L., Mutmansky, Ramani, & Wang, 1997) occurs can easily be calculated using the following expression:
Where; ° ts = surface rock temperature = 25 C
5.5 Result Using the relationship presented in chapter 5.4 the following depths could be determined 67
5.5.1 Calculations of the critical depth due to autocompression The climate in Zambia varies as in most places in seasons over the year (Weatherbase, 2013). The calculations have therefore been performed for the least favorable conditions for the atmosphere in the mine, since the ventilation system should be designed to handle the worst conditions. The least favorable conditions are likely to occur in November to December when the climate is both warm and very humid. The average high wetbulb temperature was estimated to be around 20°C ± 1°C this was based on the dewpoint of around 20°C (WeatherSpark, 2013)for these months and a high dry bulb temperature of 23°C (YR, 2013). The following critical depth could then be estimated using the same relationship:
5.5.2 Calculations of the critical rock temperature Since a slight uncertainty of the validation of the surface rock temperature was present during these calculations, the surface rock temperature was calculated for a maximum value of 26°C and a minimum value of 23°C. By doing so the critical depth came out as an interval rather than a fixed depth.
= 833 m
= 1000 m
68
5.6 Conclusions As mining reaches greater depths the role of mine ventilation is going to be ever more important for the mining operations. The heat being the major contaminant in deep mines will ultimately determine to which depth mining operations can be performed. This depth can however be extended if attention is kept to the importance of mine ventilation and its characteristics.
5.6.1 General conclusions The following list is a summary of the recommendations revealed and the conclusions made from the literature study.
Heat cannot be diluted nor can it be filtered the only solution to keep heat levels low is to identify and minimize the effects of the heat source/sources, to ventilate and/or to employ air conditioning. To prevent unnecessary heating of the fresh air mined out areas should be sealed off. The vertical infrastructure of the ventilation system should if feasible be kept separate from the mines infrastructure in order to decrease airflow resistance Dehumidification of the air makes cooling of the human body through evaporation easier. If mine workers are exposed to excessive heat for a longer period of time the productivity goes down and accident rates are increased. If the air temperature reaches higher levels than the temperature of the human body the environment can be harmful for humans. A minimum air velocity of 0.5-07 m/s ensures that most of the benefits of higher airflows are harnessed. If the air reaches temperatures of the human body; excessive ventilation needs to be avoided and air conditioning should be employed.
5.6.2 Conclusions for Mufulira and Mindola Based on the literature study, the information and experience gathered in the two case studies and the analysis presented in chapter 5.4 the following conclusions have been made for the two mines. These conclusions should not be considered as facts but as inspiration for a more thorough analysis and further investigation in the subject.
The current situation in the two mines is that the deeper parts of the mine are experiencing higher atmospheric temperatures than what is desirable. The solution which is being employed in order to solve this situation is to increase the airflow in the mines. No air-conditioning is currently employed or planned for. However based on the results from the analysis presented in chapter 5.4 the conclusion can be made that both mines are currently mining at or have already passed the critical depth. This conclusion is also confirmed by the reports from the mine of the very high registered temperatures at this depth. At this depth any increase of airflow will only add to the heat load and should therefore be kept at a practical minimum. The only solution to decrease the heat
69 load below this depth is to employ air-conditioning. The recommendation is therefore that the expansion of the ventilation system should be performed with the complementation of air- conditioning.
Both mines struggle with getting the fresh air to the production due to air leakages this is particularly a problem in Mufulira were old mined out stopes have been reopened and now acts as giant radiators which heats up the fresh air before it reaches its destination. The solution that is being employed is to seal these stopes. With reference to the analysis performed in chapter 5.4 it is considered that the change in mining method which was analyzed also would solve this problem. If the mining method would be changed to sublevel open stoping with backfill then the stopes would remain open only for a short period of time before they would be backfilled. Since the stope walls has been identified as the major heat source this would most likely result in a substantial lowering of the atmospheric temperature in the mine. To further minimize the air leakages it is recommended that automated gates are installed and that all abandoned parts of the mine are sealed off.
70
5.7 References
5.7.1 Literature Atlas Copco.(2003). Underground Mining Methods Ed. 1.Orebro: Atlas Copco. de la Vergne, J. (2003). Hard rock miner's handbook, Ed 3. Ontario, Canada: McIntosh Engineering Inc.
Hartman, H. L., Mutmansky, J. M., Ramani, R. V., & J., W. Y. (1997). Mine Ventilation and Air conditioning Ed 3. Canada: John Wiley & Sons.
Hartman, H., &Mutmansky, J. (2002). Introductory Mining Engineering -2nd ed. Hoboken, New Jersey: John Wiley and Sons, Inc.
Spalding, J. (1949). Deep Mining: An advanced Textbook for Graduates in Mining and for Practising Mining Engineers. London, UK: Mining Publications.
5.7.2 Publications and reports Wendorff, M. (2005).Sedimentary genesis and lithostratigraphy of NeoproterozoicMegabreccia from Mufulira. African Earth Science 42, 61-81.
Brake, R., &Fulker, B. (2000).The Ventilation and Refrigeration Design for Australia's Deepest and Hottest Underground Operation - the Enterprise Mine. MassMin 2000 (pp. 611-621). Brisbane, Queensland, Australia: The Australasian Institute of Mining and Metallurgy.
Pareja, L. D. (2000). Deep underground Hard-Rock Mining - Issues,Strategies and Alternatives. Kingston, Ontario, Canada: Queens University.
E K Chanda and C Katongo. Evolution of Vertical Crater Retreat Mining at Mindola Mine, Zambia, ,MassMin 2000 Brisbane, Queensland, Australia: The Australasian Institute of Mining and Metallurgy.
5.7.3 Electronic documents ZCCM.(2013, 06 10). Zambia Consolidated Copper Mines Ltd. Investments Holdings. Retrieved from http://www.zccm-ih.com.zm/
Encyclopedia Britannica academic online edition. (2013, aug 13). Retrieved from http://www.britannica.com.proxy.lib.ltu.se/EBchecked/topic/396096/Mufulira
MSHA.(2013 – 09 - 01). Hazard Alert: Heat Stress. Retrieved 09 03, 2013, from MSHA: http://www.msha.gov/Alerts/HeatStressAlert72013.pdf
71
Weatherbase.(2013 – 09 - 02). Retrieved 09 3, 2013, from http://www.weatherbase.com/weather/weatherall.php3?s=16576&units=metric
WeatherSpark.(2013 – 09 - 02). Retrieved 2013, from http://weatherspark.com/averages/29089/Lusaka-Zambia
YR. (2013 – 09 - 02). Retrieved 09 03, 2013, from http://www.yr.no/place/Zambia/Copperbelt/Kitwe/statistics.html
5.7.4 Unpublished material MOPANI.(2013a). MOPANI Corporate Brochure.Copperbelt, Zambia: Mopani.
MOPANI.(2013b). Geotechnical considerations of the Mufulira In-situ leaching project.Mufulira Rock mechanics Department.
MOPANI.(2013c). Mufulira Mine Site, Description of Mining Methods.Mufulira Mining Department.
Mopani.(2013d). NKANA SUMMARY GEOLOGY.
Mopani. (2013e). PROCEDURE FOR UNDERGROUND WORKING PLACES IN CASE OF A SEISMIC EVENT OCCURRENCE.
Mopani.(2013f). HANDBOOK OF DEVELOPMENT SUPPORT STANDARD.
5.7.5 Interviews
Mufulira: Seth Owusu Sarpong (Senior Rock mechanics Engineer at Mopani Mufulira). 2013-05-06 Mufulira Mine, Mufulira, Zambia [email protected]
MabvutoKampeni (Head of Mine Planning at Mopani Mufulira) 2013-05-07 Mufulira Mine, Mufulira, Zambia [email protected]
Mindola:
Mr Charles Katongo (Superintendent - Rock mechanic at Mopani Mindola). 2013-05-02 Mindola Mine, Kitwe, Zambia [email protected]
72
Victor Bwalya (Head of Mine Planning at Mopani Mindola) 2013-05-02 Mindola Mine, Kitwe, Zambia [email protected]
Howard Mutawa (Ventilation Superintendent at Mopani Mindola) 2013-05-03 Mindola Mine, Kitwe, Zambia [email protected]
Stanley Chasauka (Principal Geologist at Mopani Mindola) 2013-05-02 Mindola Mine, Kitwe, Zambia [email protected]
73
Appendix
Appendix A Support Standard for Mufulira a. The maximum unsupported span at Mufulira Mine Site is 5m.This is the allowable distance from the face to the nearest supported portion. b. The support pattern is the diamond pattern and the support material differs from the mining drive to the footwall drive and access cross cuts.
PRIMARY SUPPORT PATTERN FOR DRIVES • Length of Bolts = 2.4m • Number of bolts in Rings: • Ring A = 9 Bolts • Ring B = 8 Bolts • Burden = 1.0m • Bolt Spacing = 1.0m • Diameter of Hole = 35 – 40mm • Inclination of Bolts/Holes as per Sections Ring A and Ring B • Possible drill perpendicular to Bedding Planes • Number of Bolts per linear metre of x/c = 8.5m • Total length of Holes per linear metre of Cross-cut = 20.4m Figure A 1 Support standards for primary support in drives. (MOPANI, 2013c)
a
Figure A 2 Support standards for secondary support in Footwall drives. (MOPANI, 2013c)
Length of Bolts = 2.4m Number of bolts in Rings: Ring A = 8 Bolts Ring B = 7 Bolts Burden = 1.0m Bolt Spacing = 1.0m Diameter of Hole = 35 – 40mm Inclination of Bolts/Holes as per Sections Ring A and Ring B Possible drill perpendicular to Bedding Planes Number of Bolts per linear meter of x/c = 7.5m Total length of Holes per linear meter of Cross-cut = 18m
b
Figure A 3 Support standards secondary support in Mining drives. (MOPANI, Mufulira Mine Site, Description of Mining Methods)
Length of Bolts = 2.0m Number of bolts in Rings: Ring A = 8 Bolts Ring B = 7 Bolts Burden = 1.0m Bolt Spacing = 1.0m Diameter of Hole = 35 – 40mm Inclination of Bolts/Holes as per Sections Ring A and Ring B Possible drill perpendicular to Bedding Planes Number of Bolts per linear meter of x/c = 7.5m Total length of Holes per linear meter of Cross-cut = 15m Access Cross cut North
c
PRIMARY SUPPORT PATTERN FOR CROSS-CUTS • Length of Bolts = 2.4m • Number of bolts in Rings: • Ring A = 9 Bolts • Ring B = 8 Bolts • Burden = 1.0m • Bolt Spacing = 1.0m • Diameter of Hole = 35 – 40mm • Inclination of Bolts/Holes as per Sections Ring A and Ring B • Possible drill perpendicular to Bedding Planes • Number of Bolts per linear metre of x/c = 8.5m • Total length of Holes per linear metre of Cross-cut = 20.4m Figure A 4 Support standards for primary support in cross-cuts. (MOPANI, 2013c)
d
Figure A 5Support standards for Junctions. (MOPANI, 2013c)
e
Figure A 6 Illustration on how mesh should be installed. (MOPANI, Mufulira Mine Site, Description of Mining Methods)
Collaring of an end should always be done when the support is within 5m past the collaring distance and wire mesh installed 10m behind the collaring point.
Meshing The wire mesh should be rolled out for installation. All mesh joints should overlap by at least 20cm and be securely fastened. Holes must not be cut in the mesh for the purpose of bleeding rock without authorization from Rock mechanics. If the mesh begins bulging excessively, it may be held back by installing another rook bolt.
f
Figure A 7 Illustration on how lacing should be installed. (MOPANI, Mufulira Mine Site, Description of Mining Methods)
Lacing Rope and Chain links used for lacing should be in good condition. The tensile strength of the lacing rope should be greater than 5 Tonnes. Lacing should follow a diamond pattern. In high stress areas, the pattern can be intensified by using additional straight strands, When two strands of lacing rope intersect, particularly passing through rook bolt loops, they should be intertwined. Lacing ropes should be joined by two chain link clamps on both sides of the loop as shown in the figure(s) below Intermediate clamps should be installed every 5.0m along the rope. This prevents more than 5m of slack in the event of failure. Tension lacing using a wire winch, to approximately 1.6 Tonnes Use clamps to keep the lacing rope tight to the rock wall where bolts protrude to far g
Figure A 8 Cable Support in the Footwall Drive. (MOPANI, 2013c)
Spacing of rings in cable bolt support should be 2m apart and the number of bolts per ring will be influenced by the width of the drive.
Figure A 9 Cable Support in the Hanging wall. (MOPANI, Mufulira Mine Site, Description of Mining Methods)
Spacing of rings should be 2m and the cable bolt holes should effectively be drilled 5m beyond the hanging wall contact. The number of bolts per ring differs from one section to the other depending on the inclination of the ore body and it will be recommended by rock Mechanics section. The choice of support for the hanging wall will be dictated by the inclination of the ore body and this should come as a recommendation from Rock Mechanics Section.
h
Figure A 10Installation of roof bolts best operating practice. ( Information receiveed at the interview with Seth Owusu Sarpong (Senior Rock mechanics Engineer at Mopani Mufulira))
i
Figure A 81 Installation of roof bolts best operating practice. (Information received at the interview with Seth Owusu Sarpong (Senior Rock mechanics Engineer at Mopani Mufulira))
j
Figure A 12 Installation of roof bolts best operating practice. (Information received at the interview with Seth OwusuSarpong (Senior Rock mechanics Engineer at Mopani Mufulira))
k
Figure A 13 Installation of roof bolts best operation practice. (Information received at the interview with Seth OwusuSarpong (Senior Rock mechanics Engineer at Mopani Mufulira))
l
Appendix B
m
n
o p
q
r
s
t
Appendix C
Table C 1 Hanging wall properties Hanging wall properties Initial element loading: Field stress and body force Unit weight: 0.027 MN/m3 Elastic type: Isotropic Poisson’s ratio: 0.2 Young’s modulus: 55500 MPa Failure Criterion: Hoek-Brown Material type: Elastic Intact compressive strength: 154 MPa mb Parameter (peak) 11.562 S Parameter (peak) 0.1353
Table C 2 Foot wall properties Foot wall properties Initial element loading: Field stress and body force Unit weight: 0.027 MN/m3 Elastic type: Isotropic Poisson’s ratio: 0.3 Young’s modulus: 68750 MPa Failure Criterion: Hoek-Brown Material type: Elastic Intact compressive strength: 183 MPa mb Parameter (peak) 10.147 S Parameter (peak) 0.1211
Table C 3 Ore Shale properties Ore shale properties Initial element loading: Field stress and body force Unit weight: 0.027 MN/m3 Elastic type: Isotropic Poisson’s ratio: 0.2 Young’s modulus: 51600 MPa Failure Criterion: Hoek-Brown Material type: Elastic Intact compressive strength: 109 MPa mb Parameter (peak) 2.375 S Parameter (peak) 0.0229
u
Table C 4 Rock fill properties Rock fill properties Initial element loading: Body force Unit weight: 0.0165 MN/m3 Elastic type: Isotropic Poisson’s ratio: 0.25 Young’s modulus: 309 MPa Failure Criterion: Hoek-Brown Material type: Plastic Intact compressive strength: 0.6 MPa mb Parameter (peak) 0.191 S Parameter (peak) 6.3e-006 Dilation Parameter 1 mb Parameter (residual) 0.191 S Parameter (residual) 6.3e-006
v
Appendix D This appendix contains the old mining sequence which left behind crown pillars. It also includes the stress measurements for the three targeted drifts during excavation of the stopes. The results act as a basecase for the other assays.
1735m
1803m
1853m
Figure D 1 Excavation step 1 with footwall drifts placed 50m from the stope
w
1735m
1803m
1853m
Figure D 2 Excavation step 2 with footwall drifts placed 50m from the stope
1735m
1803m
1853m
Figure D 3 Excavation step 3 with footwall drifts placed 50m from the stope
x
1735m
1803m
1853m
Figure D 4 Excavation step 4 with footwall drifts placed 50m from the stope
1735m
1803m
1853m
Figure D 5 Excavation step 5 with footwall drifts placed 50m from the stope
y
1735m
1803m
1853m
Figure D 6 Excavation step 6 with footwall drifts placed 50m from the stope
1735m
1803m
1853m
Figure D 7 Excavation step 7 with footwall drifts placed 50m from the stope
z
1735m
1803m
1853m
Figure D 8 Excavation step 8 with footwall drifts placed 50m from the stope
1735m
1803m
1853m
Figure D 9 Excavation step 9 with footwall drifts placed 50m from the stope
å
Appendix E This appendix contains the mining sequence with backfill which does not leave behind crown pillars. In this model the footwall drift is placed at a distance of 50m from the stopes. It also includes the stress measurements for the three targeted drifts during excavation of the stopes.
1735m
1803m
1853m
Figure E 1 Excavation step 1 with footwall drifts placed 50m from the stope
ä
1735m
1803m
1853m
Figure E 2 Excavation step 2 with footwall drifts placed 50m from the stope
1735m
1803m
1853m
Figure E 3 Excavation step 3 with footwall drifts placed 50m from the stope
ö
1735m
1803m
1853m
Figure E 4 Excavation step 4 with footwall drifts placed 50m from the stope
1735m
1803m
1853m
Figure E 5 Excavation step 5 with footwall drifts placed 50m from the stope
aa
1735m
1803m
1853m
Figure E 6 Excavation step 6 with footwall drifts placed 50m from the stope
1735m
1803m
1853m
Figure E 7 Excavation step 7 with footwall drifts placed 50m from the stope
bb
1735m
1803m
1853m
Figure E 8 Excavation step 8 with footwall drifts placed 50m from the stope
cc
Appendix F This appendix illustrates the mining sequence with backfill which does not leave behind crown pillars. In this model the footwall drift is placed at a distance of 30m from the stopes. It also includes the stress measurements for the three targeted drifts during excavation of the stopes.
1735m
1803m
1853m
Figure F 1 Excavation step 1 with footwall drifts placed 30m from the stope
dd
1735m
1803m
1853m
Figure F 2 Excavation step 2 with footwall drifts placed 30m from the stope
1735m
1803m
1853m
Figure F 3 Excavation step 3 with footwall drifts placed 30m from the stope
ee
1735m
1803m
1853m
Figure F 4 Excavation step 4 with footwall drifts placed 30m from the stope
1735m
1803m
1853m
Figure F 5 Excavation step 5 with footwall drifts placed 30m from the stope
ff
1735m
1803m
1853m
Figure F 6 Excavation step 6 with footwall drifts placed 30m from the stope
1735m
1803m
1853m
Figure F 7 Excavation step 7 with footwall drifts placed 30m from the stope
gg
1735m
1803m
1853m
Figure F 8 Excavation step 8 with footwall drifts placed 30m from the stope
hh