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New ventilation design criteria for underground metal mines based upon the "life-cycle" airflow demand… Kocsis, Karoly-Charles P. 2009

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NEW VENTILATION DESIGN CRITERIA FOR UNDERGROUND METAL MINES BASED UPON THE “LIFE-CYCLE” AIRFLOW DEMAND SCHEDULE  by  Karoly-Charles Kocsis P. Eng. Engineering Degree, Mining University of Petrosani, Romania, 1987 M.A.Sc., Laurentian University of Sudbury, Canada, 1998  A THESIS SUBMITTED IN PARTIAL FULFILLMENT OF THE REQUIREMENTS FOR THE DEGREE OF DOCTOR OF PHILOSOPHY  in The Faculty of Graduate Studies (Mining Engineering)  THE UNIVERSITY OF BRITISH COLUMBIA (Vancouver) August 2009  © Karoly-Charles Kocsis, 2009  ABSTRACT An underground ventilation system which requires electricity to operate the fans and fossil fuel for heating in winter, can account for a significant portion of a mine’s energy consumption. Recent studies have shown that the ventilation system in highly mechanized metal mines could be responsible for 40-60% of the mines’ energy consumption. Presently, underground ventilation systems are designed more towards the “worst-case-scenario” with respect to airflow demand, which usually occurs well into the future of a mine’s operating life. As a result, within the early stages of operation, the mine’s intake air volume could be well in excess of its “true” ventilation needs. Such ventilation systems are inefficient and wasteful and this design approach needs to change if Canadian mines are to remain competitive while attempting to reduce their carbon footprints. The concept of setting up “smart”, activity based underground ventilation systems is a way to address such redundant use of air in mines. In simple terms this would entail monitoring the production areas for the presence of activity and mining equipment, based upon which, appropriate air volumes would be delivered accordingly. However, determining the energy consumption of such ventilation systems and their cost benefits is challenging using conventional analysis techniques. Probably the most important task in optimizing the ventilation system of an existing mine is to accurately determine its “true” intake air volume requirement, in order to identify and determine the ventilation redundancy throughout the mine. Two conventionally analyzed case studies presented in this thesis highlight the complexity and difficulty in determining the potential ventilation redundancy in large and deep multilevel metal mines. Challenges include gaining adequate data to assess the dynamic nature of the production activities that continually redefine where ventilation is required. This thesis introduces a new ventilation design concept where the “true” intake air volume requirement for the life of a mine is determined by means of mining process simulations performed on a model developed using AutoModTM. Based upon the output data generated by discrete-event process simulations, the “life-cycle” airflow demand schedule of a mine was determined for “traditional” versus “activity-based” ventilation practices. Furthermore, based upon the “life-cycle” airflow demand schedule, the ventilation system was solved and balanced using ventilation simulation. The output data generated by ventilation simulation was used to determine the economic and environmental benefits of an activity based ventilation system versus a traditional one. KEYWORDS: mine ventilation, ventilation system optimization, discrete-event process simulation, ventilation simulation, life-cycle airflow demand schedule, ventilation on demand, activity based ventilation controls, AutoModTM, energy savings, GHG emissions reductions.  ii  TABLE OF CONTENTS ABSTRACT............................................................................................................................. ii LIST OF TABLES ................................................................................................................ vii LIST OF FIGURES .................................................................................................................x 1.  INTRODUCTION...........................................................................................................1 1.1 Ventilation Challenges in Deep Mines ....................................................................1 1.1.1 The Depth Challenge ......................................................................................1 1.1.2 The Effect of Increased Mechanization ..........................................................2 1.1.3 Health and Environmental Considerations .....................................................3 1.2 Statement of the Problem.........................................................................................4 1.3 Thesis Objectives .....................................................................................................4 1.4 Limitations of this Research ....................................................................................5 1.5 Organization of this Thesis ......................................................................................5  2.  LITERATURE REVIEW ..............................................................................................7 2.1 Underground Ventilation Systems – Background Information ...............................7 2.2 Review and Quantification of the Current Ventilation Design Criteria in Canadian Mines........................................................................................................................9 2.2.1 Heat Sources in Underground Mines............................................................10 Auto-compression .....................................................................................................11 Mining Equipment ....................................................................................................11 Strata (geothermal gradient) .....................................................................................12 Explosives .................................................................................................................12 2.2.2 Health and Safety Concerns in Canadian Mines - Exposure Limits and Guidelines .....................................................................................................12 2.3 Ventilation Network Analysis – Background Information....................................14 2.3.1 Methods of Solving Mine Ventilation Networks..........................................15 2.3.2 Ventilation Network Simulation Programs...................................................17 2.4 Ventilation Control Systems ..................................................................................18 2.4.1 Examples of Event-Based Ventilation Control Systems ..............................21 Example 1: Event-Based Ventilation on Demand System at Vale Inco’s Creighton Mine, Sudbury, Canada.................................................................................22 Example 2: Equipment Tracking and Ventilation Control System at Barrick’s Goldstrike Mine, Nevada, USA ....................................................................24 Example 3: Ventilation on Demand System at LKAB’s Malmberget Iron Ore Mine, Sweden..........................................................................................................28 Example 4: Real-Time Airflow Monitoring System at Capcoal Central Colliery and Cannington Mine, Queensland, Australia.....................................................30 2.4.2 Example of a Real-Time Ventilation Control System ..................................32 Example 5: Bestech’s Real-Time Ventilation Control System (MCS) ...................32 2.5 Summary ................................................................................................................34 iii  3.  METHODOLOGY .......................................................................................................35 3.1 Underground Operations Survey ...........................................................................36 3.2 Ventilation Redundancy and its Importance in Measuring the Efficiency of Underground Ventilation Systems.........................................................................36 3.3 Mining Process Model Development Using AutoModTM .....................................41 3.4 Ventilation System Design Based upon the Mine’s Life-Cycle Airflow Demand Schedule.................................................................................................................42 3.5 Economic and Environmental Benefits .................................................................42  4.  UNDERGROUND OPERATIONS SURVEY............................................................43 4.1 Mining Methods Employed by Canadian Mines ...................................................43 4.2 Identification of Canadian Mines that Would Benefit the Most from an Activity Based Ventilation Control System.........................................................................44 4.2.1 Hardrock Mines – Base and Precious Metal Mines......................................45 4.2.2 Softrock Mines..............................................................................................46  5.  Case Studies: COMPARISON OF CURRENT VENTILATION DESIGN CRITERIA WITH THE NEW DESIGN CONCEPT IN DEEP METAL MINES 48 5.1 Introduction............................................................................................................48 5.2 Ventilation Redundancy and its Importance in Measuring the Efficiency of Underground Ventilation Systems.........................................................................49 5.3 Case Study #1: Determination of Ventilation Redundancy at Creighton Mine – Vale Inco................................................................................................................49 5.3.1 Air Volume Requirements at Creighton Mine..............................................50 5.3.2 Mine Activity Monitoring Through Temperature Studies............................52 5.3.3 Production Blast Monitoring.........................................................................57 5.3.4 Determination of Ventilation Redundancy at Creighton Mine.....................60 5.3.5 Prediction of Cost Savings through the Elimination of Ventilation Redundancy at Creighton Mine ....................................................................61 5.4 Case Study #2: Ventilation Utilization Efficiency at Kidd Creek Mine – Xstrata PLC ........................................................................................................................63 5.4.1 First Stage Review - Brainstorming Forum ..................................................64 5.4.2 Second Stage Review - Future Production Planning Based Analysis ..........64 5.4.3 Third Stage Review - Retrospective Activity Based Analysis based upon the Mine’s Database (SIMS)...............................................................................68 5.4.4 Variance between “Future Production Planning” and “Retrospective Activity Based” Methods of Analysis...........................................................74 5.4.5 Ventilation Redundancy at the Kidd Creek Mine - Discussions ..................75 5.5 Challenges and Difficulties in Determining Ventilation Redundancy in Large and Deep Multilevel Operations...................................................................................77 5.6 The Need to Change Current Ventilation Design Practices...................................78  6.  MINING PROCESS MODEL DEVELOPMENT USING AutoModTM SIMULATOR................................................................................................................81 6.1 Development of the “Mining Process” Model.......................................................81 iv  6.1.1 Mining Processes and Systems .....................................................................81 6.1.2 Ventilation System........................................................................................86 6.1.3 Input Data......................................................................................................89 6.1.4 Output Data Manipulation and Analysis.......................................................93 6.1.5 Summary - Output Data Manipulation .......................................................109 7.  DETERMINATION OF SHORT-TERM AND LIFE-CYCLE AIRFLOW REQUIREMENTS FOR METAL MINES ..............................................................111 7.1 Determination of Short-Term Airflow Requirements..........................................111 7.1.1 Short-Term Airflow Requirements – Traditional Ventilation Practice ......112 7.1.2 Short-Term Airflow Requirements – Activity Based Ventilation ..............117 7.2 Determination of Life-Cycle Airflow Requirements...........................................119 7.2.1 Life-Cycle Airflow Requirements – Traditional Ventilation Practice........120 7.2.2 Life-Cycle Airflow Requirements – Activity Based Ventilation ...............122  8.  VENTILATION SYSTEM DESIGN BASED UPON THE LIFE-CYCLE AIRFLOW DEMAND SCHEDULE .........................................................................125 8.1 Solving and Balancing the Primary and Auxiliary Ventilation Systems of “Mine A” for Years 2008–2010 and for Traditional Air Volume Requirements ...........127 8.1.1 Ventilation Modelling for Years 2008–2010 (Traditional Ventilation) .....127 8.2 Solving and Balancing the Primary and Auxiliary Ventilation Systems of “Mine A” for Years 2008-2010 and Activity Based Air Volume Requirements ...........138 8.2.1 Ventilation Modelling for Years 2008–2010 (Activity Based Ventilation) 139 8.3 Power Consumption and Operating Cost Comparison – Traditional versus Activity Based Ventilation Practices ...................................................................146  9.  ECONOMIC AND ENVIRONMENTAL BENEFITS OF THIS NEW VENTILATION DESIGN APPROACH ..................................................................151 9.1 Cost Estimate of an Event-Based Ventilation Control System............................151 9.2 Cost Estimate for a New Intake System at “Mine A”..........................................153 9.3 Environmental Benefits of this New Ventilation Design Approach....................166 9.3.1 Green House Gas Emissions – Background Information ...........................166 9.3.2 Projected Reductions in GHG Emissions at “Mine A” Generated by an Activity Based Ventilation Control System................................................167  10.  THESIS SUMMARY – RESULTS AND DISCUSSIONS ......................................169  11.  CONCLUSIONS .........................................................................................................175  12.  THESIS CONTRIBUTION AND FUTURE WORK ..............................................178 12.1 Thesis Contributions ............................................................................................178 12.2 Future Work .........................................................................................................179  BIBLIOGRAPHY ................................................................................................................180 APPENDICES ......................................................................................................................187 v  Appendix A ...........................................................................................................................187 Appendix B ...........................................................................................................................195 B.1 Solving and Balancing the Primary and Auxiliary Ventilation Systems of “Mine A” for Years 2011–2013, Years 2014-2017, Years 2018-2020 with Traditional Air Volume Requirements ...................................................................................195 B.1.1 Ventilation Modelling for Years 2011–2013 (Traditional Ventilation) .....195 B.1.2 Ventilation Modelling for Years 2014–2017 (Traditional Ventilation) .....200 B.1.3 Ventilation Modelling for Years 2018–2020 (Traditional Ventilation) .....204 B.2 Solving and Balancing the Primary and Auxiliary Ventilation Systems of “Mine A” for Years 2011–2013, Years 2014-2017, Years 2018-2020 with Activity Based Air Volume Requirements ...................................................................................207 B.2.1 Ventilation Modelling for Years 2011–2013 (Activity Based Ventilation) 207 B.2.2 Ventilation Modelling for Years 2014–2017 (Activity Based Ventilation) 211 B.2.3 Ventilation Modelling for Years 2018–2020 (Activity Based Ventilation) 214  vi  LIST OF TABLES Table 1:  Ventilation Requirements in Canadian Mines Based on Diesel Engine Certification [Gangal & Grenier, 2002]................................................................ 10  Table 2:  Proposed/Adopted Exposure Limits and Guidelines within the Mining Industry [Gangal & Grenier, 2002] ..................................................................................... 14  Table 3:  Classification of the Canadian mines based upon mining method utilization frequency............................................................................................................... 45  Table 4:  Temperature and air volume requirements as a function of mining depth at Creighton Mine - Vale Inco [Stachulak, 1989]..................................................... 51  Table 5:  Increasing airflow and power needs at Creighton Deep [Hardcastle & Kocsis 2004] ..................................................................................................................... 52  Table 6:  Combined ventilation operating costs/savings with time-of-day and activity based controls [Hardcastle et al, 2006] ........................................................................... 61  Table 7:  Primary ventilation system operating cost/savings with time-of-day controls [Hardcastle et al, 2006] ......................................................................................... 62  Table 8:  Auxiliary ventilation operating costs/savings with time-of-day and activity based ventilation controls [Hardcastle et al, 2006] ......................................................... 62  Table 9:  An example of airflow allocations per level (in m3/s) and for all the lower mine based on the 18-month production plan before and after allowances for leakage and to avoid recirculation (shaded areas represent air volume substitutions for production activities, to provide minimum airflow velocities or a to prevent uncontrolled recirculation within the auxiliary system) [Hardcastle, Kocsis & Li, 2006] ..................................................................................................................... 67  Table 10: Sample of the lower mine airflow requirements per level (in m3/s) by “month” derived from historical equipment deployment data (shaded areas denote substitutions for minimum airflow or required airflow to prevent recirculation within the auxiliary system) [Hardcastle, Kocsis & Li, 2006] ............................. 70 Table 11: Sample of the lower mine airflow requirements per level (in m3/s) by “week” and “day” derived from historical equipment deployment data (shaded areas denote substitutions for minimum airflow or required airflow to prevent recirculation within the auxiliary system) [Hardcastle, Kocsis & Li, 2006] ............................. 71 Table 12: Fan operating duties and power requirements for primary system (years 2008 – 2010) – Traditional ventilation........................................................................... 132 vii  Table 13: Auxiliary fans operating duty and power requirements within the mining block Traditional ventilation ........................................................................................ 137 Table 14: Fan operating duties and power requirements for primary system (years 2008 2010) – Activity based ventilation....................................................................... 141 Table 15: Auxiliary fans operating duty and power requirements within the mining block – Activity based ventilation practice...................................................................... 144 Table 16: Intake Airflow, Power Consumption and Operating Costs - Traditional Ventilation Practice............................................................................................. 147 Table 17: Intake Airflow, Power Consumption and Operating Costs - Activity Based Ventilation Practice............................................................................................. 147 Table 18: Power Consumption & Operating Cost Savings – Activity Based vs. Traditional Ventilation........................................................................................................... 147 Table 19: Fan power in the primary system with new intake infrastructure (Years 2014 2017) – Traditional ventilation practice............................................................. 158 Table 20: Fan power in the primary system with new intake infrastructure (years 2018 2020) – Traditional ventilation practice............................................................. 159 Table 21: Ventilation Operating & Capital Costs for Activity Based vs. Traditional Ventilation – Without New Intake Infrastructure ............................................... 161 Table 22: Ventilation Operating & Capital Costs for Activity Based vs. Traditional Ventilation – With New Intake Infrastructure .................................................... 161 Table 23: GHG Emissions in CO2 Equivalent (Mt) by Industry Sub-Sectors [Canada’s Emissions Outlook, 1999]................................................................................... 167 Table 24: Reductions in GHG emissions generated by ventilation control system at “Mine A”........................................................................................................................ 168 Table 25: Mine A’s life-cycle airflow demands for traditional & activity based ventilation ............................................................................................................................. 172 Table 26: Mine A’s ventilation operating costs for traditional and activity based ventilation ............................................................................................................................. 172 Table 27: Fan operating duties and power requirements for primary system (years 2011 2013) – Traditional ventilation........................................................................... 198  viii  Table 28: Fan Operating Duties and Power Requirements for Primary System (Years 2014 2017) – Traditional ventilation........................................................................... 203 Table 29: Fan Operating Duties and Power Requirements for Primary System (Years 2018 2020) – Traditional ventilation........................................................................... 206 Table 30: Fan Operating Duties and Power Requirements for Primary System (Years 2011 2013) – Activity based ventilation....................................................................... 210 Table 31: Fan operating duties and power requirements for primary system (years 2014 2017) – Activity based ventilation....................................................................... 213 Table 32: Fan operating duties and power requirements for primary system (years 2018 2020) – Activity based ventilation....................................................................... 217  ix  LIST OF FIGURES Figure 1: Traditional Ventilation Design Practice [Kocsis et al, 2003] ............................... 18 Figure 2: Generic Ventilation Control System Components [Kocsis et al, 2004] ............... 20 Figure 3: Creighton mine’s ventilation on demand schematic [O’Connor, 2008] ............... 23 Figure 4: Barrick’s Equipment Tracking and Ventilation Control System.......................... 25 Figure 5: Block diagram of model control system components [Tonnos et al, 2008] ......... 32 Figure 6: Location of temperature and humidity monitoring sensors on 7400L – mucking operation [Hardcastle & Kocsis, 2004]................................................................. 54 Figure 7: Typical dry-bulb temperatures plot around a 250 kW LHD moving ore from a draw-point to the orepass [Hardcastle & Kocsis, 2004] ....................................... 55 Figure 9: Activity log derived from temperature monitoring at a mucking draw-point operation [Hardcastle & Kocsis, 2004]................................................................. 58 Figure 10: Sample of blast contaminant decay monitoring [Hardcastle & Kocsis, 2004] ..... 59 Figure 11: Comparison of the lower mine’s ventilation requirements derived from historic production data (daily, weekly, monthly) and from future monthly production plans [Hardcastle, Kocsis & Li, 2006].................................................................. 72 Figure 12: Comparison of the number of active levels within the lower mine derived from historic production data (daily, weekly, monthly) and from future monthly production plans [Hardcastle, Kocsis & Li, 2006] ............................................... 73 Figure 13: Representation of a mining block with 2 production stopes (VRM 1 &VRM 2), 2 developments, 2 ore & rock passes....................................................................... 83 Figure 14: Locations of the auxiliary fans in this multi-level mining block – “Mining Process” model...................................................................................................... 85 Figure 15: Spreadsheet based mine activity input data utility for the “Mining Process” model (i.e. Activity Breaks, VRM1)................................................................................ 90 Figure 16: Spreadsheet based mine activity input data utility for the “Mining Process” model (Development 1, Development 2)......................................................................... 91 Figure 17: Fan volumes data entry using MSExcelTM worksheet .......................................... 92 x  Figure 18: Custom report showing when and for how long each auxiliary fan is operating in hours (absolute clock) ........................................................................................... 95 Figure 19: “Fan Summary” macro developed for the “Mining Process” model.................... 96 Figure 20: VRM 1 production activity durations and occurrence times (in AC)................... 98 Figure 21: VRM 2 production activity durations and occurrence times (in AC)................... 99 Figure 22: Development 1 and Development 2 activity durations and occurrence times (in AC)...................................................................................................................... 100 Figure 23: Fan activity start and end times in hours (absolute clock) for VRM 1 and VRM 2 ............................................................................................................................. 101 Figure 24: Fan activity start and end times in hours (absolute clock) for Development 1 and Development 2 .................................................................................................... 102 Figure 25: Air volume required at the mining block (with simulation time divided into 30minute intervals) ................................................................................................. 105 Figure 26: Air volume required at the mining block (with simulation time divided into 30minute intervals and early fan start).................................................................... 106 Figure 27: Air volume required at the mining block (with simulation time divided into 5minute intervals) ................................................................................................. 107 Figure 28: Air volume required at the mining block (with simulation time divided into 5minute intervals and early fan starts) .................................................................. 108 Figure 29: Airflow distribution within the mining block – traditional ventilation practice . 115 Figure 30: Airflow requirements for production and secondary activities – Year 2008 (Traditional Ventilation Practice) ....................................................................... 116 Figure 31: Airflow requirements for production and secondary activities – Year 2008 (Activity Based) .................................................................................................. 118 Figure 32: Forecasted production requirements for the remaining operating life of “Mine A” ............................................................................................................................. 121 Figure 33: Life-Cycle Airflow demand Schedule for “Mine A” – Traditional vs. Activity Based Ventilation Practice.................................................................................. 124  xi  Figure 34: Airflow distribution and fan operating duties in the primary system for years 2008 – 2010 - Traditional ventilation.......................................................................... 130 Figure 35: Airflow distribution and fan duties within the mining block’s auxiliary ventilation system - Traditional ventilation.......................................................................... 135 Figure 36: Airflow distribution and fan operating duties in the primary system for years 2008 – 2010 - Activity based ventilation...................................................................... 140 Figure 37: Airflow distribution & fan operating duties within the mining block’s auxiliary ventilation system - Activity based ventilation ................................................... 143 Figure 38: Mine A’s intake air volume for “traditional” versus “activity based” ventilation ............................................................................................................................. 148 Figure 39: Mine A’s installed fan power for “traditional” versus “activity based” ventilation ............................................................................................................................. 149 Figure 40: Mine A’s operating costs for “traditional” versus “activity based” ventilation .. 149 Figure 41: Savings in operating costs generated by an “activity based” ventilation system versus “traditional” ventilation at “Mine A” ...................................................... 150 Figure 42: Airflow distribution in the primary ventilation system with new intake system for years 2014 - 2017 - Traditional ventilation ........................................................ 155 Figure 43: Airflow distribution in the primary ventilation system with new intake system for years 2018 - 2020 - Traditional ventilation ........................................................ 156 Figure 44: Combined capital and operating costs for “traditional” versus “activity based” ventilation at “Mine A”....................................................................................... 165 Figure 45: Airflow distribution and fan operating duties in the primary system for years 2011 – 2013 - Traditional ventilation.......................................................................... 197 Figure 46: Airflow distribution and fan operating duties in the primary system for years 2014 – 2017 - Traditional ventilation.......................................................................... 201 Figure 47: Airflow distribution and fan operating duties in the primary system for years 2018 – 2020 - Traditional ventilation.......................................................................... 205 Figure 48: Airflow distribution and fan operating duties in the primary system for years 2011 – 2013 - Activity based ventilation ..................................................................... 209 Figure 49: Airflow distribution and fan operating duties in the primary system for years 2014 - 2017 - Activity based ventilation ...................................................................... 212 xii  Figure 50: Airflow distribution and fan operating duties in the primary system for years 2018 - 2020 - Activity based ventilation ........................................................................................ 216  xiii  ACKNOWLEDGEMENTS I would like to thank the management and staff of Vale Inco Limited (Sudbury Operations) and Xstrata PLC (Kidd Creek Mine) for their support and for permitting access to the underground operations and mining activity data for this research work. I would like to express my profound thanks to Dr. Stephen Hardcastle and Gary Li from CANMET - Mining and Mineral Sciences Laboratories, Dr. Yassiah Bissiri and Dr. Gregory Baiden from Penguin Automated Systems Inc., for their valuable contribution, guidance and support for the development of mining process models and ventilation modelling tools without which this research work would not have been possible. I am deeply grateful to my thesis supervisor Dr. Robert Hall, for his valuable support, supervision and encouragement through the never ending process of preparing my thesis. His moral support and guidance enabled me to complete this research work. Sincerely thanks to the Supervisory Committee associated to this program, Dr. Malcolm Scoble, Dr. Scott Dunbar and Dr. Michael Hitch for their guidance and valuable time allocated to review this thesis. I would like to acknowledge the financial support of CANMET-Mining and Mineral Sciences Laboratories. In particular I would like to thank Michel Grenier, Louise Laverdure, my co-workers and administration staff from the Sudbury Laboratory for their continual support and encouragement. Finally, I would like to thank my wife, Eniko and my children, Zsolt and Monika, for putting up with me during this time period. This work was way too often placed ahead of our family matters and events.  xiv  1.  INTRODUCTION  1.1  Ventilation Challenges in Deep Mines  Worldwide, metal mines are going deeper. For example, in Canada there are six mechanized metal mines that are planning extraction of ore reserves at 3,000m depth. Working at such depth challenges all aspect of mining technically and economically, including the delivery of fresh air to the production areas. With respect to the Canadian metal mines, three general trends can be observed [Hardcastle et al, 2005]: •  Mines are getting deeper  •  Mining operations are becoming increasingly mechanized, and  •  Health and environment standards for the underground workers are becoming more stringent  To varying degrees, these observed trends coupled with increasing energy costs continue to challenge the economic sustainability of the mines, including the economic provision of ventilation in large and deep metal mines.  1.1.1 The Depth Challenge Depth can affect the ventilation economics in four ways. Firstly, as the distance from surface to the production workings increases, the associated ventilation capital and operating costs will increase linearly. This can be demonstrated through a basic airflow relationship. For a given airway this airflow relationship can be defined as: P = RQ2  (Pa)  (1)  The parameter R (kg/m7) is the resistance of the airway and it is the factor that governs the amount of airflow Q (m3/s), which will pass when this airway is subjected to a given pressure differential P (Pa). The frictional pressure drop (P) along this airway is a function of its length. This is because the resistance of the airway (R) is defined by:  LC (kg/m7) (2) A3 Where k = friction factor (kg/m3), L = airway length (m), C = airway perimeter (m) and A = R=k  cross sectional area (m2). Furthermore, the fan power needed to overcome the frictional pressure losses along the airway can be defined as:  FP = PQ = RQ3 (W)  (3) 1  Equation (3) shows that the fan power (FP) is linearly proportional to the resistance of the airway (R), hence the length of the airway (L). But more importantly this equation shows that there is a cubic relationship between the fan power (FP) and the passing airflow (Q). Secondly, depth can affect the ventilation economics due to increased leakage as the airways become older and more production levels are added to the mine’s existing infrastructure. For example, as the development and production levels deepen, an addition of 10% to a mine’s total intake air volume is needed to compensate for leakage could increase the ventilation system’s operating cost by 30-35%. This shows the need to optimize and manage the intake fresh air delivered underground. Thirdly, as more intake air volume is needed to compensate for increased airflow leakage within the primary ventilation system, the pressure requirements from the primary fans and consequently the pressure differentials between the intake and exhaust airways would also increase. As a result, airflow leakage within the mines would further increase. Fourthly, and most importantly, is that as production workings deepen, the air temperature in a mine will increase as a result of auto-compression of the air down the intake airways and due to the heat transferred from the surrounding strata. Unfortunately, this is unavoidable unless ameliorative measures are taken. To diminish this heat gain and to maintain the same cooling capacity of the air for the removal of machine heat, the intake air volume in the mines needs to increase with depth. For example, in a Canadian mine already operating at 2,000m below surface, an additional 300m depth can increase the mine’s intake air volume requirement by 20%. Due to the cubic relationship between the mine’s supplied air power and the airflow, the ventilation system’s energy consumption and consequently its operating cost can increase by 60-70% [Hardcastle et al, 2005]. This clearly shows that with increasing depth any increase in the mines’ intake air volume to combat the additional heat transferred to the airflow from auto-compression and strata can dramatically increase the ventilation system’s energy consumption and consequently its operating costs.  1.1.2 The Effect of Increased Mechanization When diesel equipment started to be introduced in Canadian mines in the 1960s, the mining equipment had diesel engines of less than 75 kW (<100 hp), and they where few in use. Since that time their number and engine size have significantly increased. Initially, the growth was in the primary haulage fleet, such as load-haul-dump (LHD) vehicles and haulage trucks. 2  However, a more recent trend is that to ensure workforce mobility, the majority of underground workers now have access to a personnel carrier. Within the haulage fleet, the size of the mining equipment and their diesel engines continue to increase. In Canada, mines now routinely employ 9.0 m3 (12 yd3) bucket capacity LHDs with 260 kW (350 hp) diesel engines. However, these are not the largest vehicles employed underground. Large metal mines in Canada now employ haulage trucks with 490 kW (650 hp) diesel engines [Hardcastle et al, 2005]. For personnel carriers and service vehicles, Ontario Ministry of Labour data (MOL, 1996 & 2003) through five surveys shows that their relative population within the underground operations has consistently increased, namely from 39% in 1977, through to 40%, 59%, 63% and most recently to 71% in 2003. In conjunction with their increasing number, between 1996 and 2003, the mine personnel vehicles contribution to the Ontario mines’ total installed diesel power has increased from 40.0% to 50.2%. Consequently, personnel vehicles are now not only the most common type of underground vehicles; they also represent more than half the mines’ installed diesel power [Hardcastle et al, 2005]. In highly mechanized metal mines the air volume requirements are usually based upon diesel exhaust dilution criteria such as 0.063 m3/s per kW of rated engine power (100 cfm/bhp) [Gangal, M. & Grenier, M., 2002]. Therefore, similar to increasing mining depths, it can be seen that increased mechanization and workforce mobility has required mines to progressively deliver more fresh air underground. Again, increasing a mine’s intake air volume can significantly increase its ventilation power requirements and operating costs.  1.1.3 Health and Environmental Considerations In underground mines ventilation is required to dilute and remove the contaminants generated by the mining operations in order to provide adequate working conditions for workers and mining equipment. The most common contaminants are: •  Gases liberated from strata, which include methane (flammable) and radioactive gasses such as radon  •  Gases generated by mining operations and mining equipment, which include the byproducts of combustion from diesel engines and fumes from explosives  •  Mineral dust liberated during blasting and throughout the ore/waste transportation and crushing processes 3  1.2  Statement of the Problem  The mining trends presented in section 1.1, including increasing depth and mechanization and more stringent health regulations would generally dictate that more air volume is required in order to provide adequate working conditions underground. However, any increase in the mine’s intake airflow can produce a disproportionate increase in power consumption and hence operating costs due to the cubic relationship between the supplied air power and the airflow. Currently, ventilation systems in metal mines are designed upon peak air volume demands based upon diesel exhaust dilution or heat criteria and are usually operated at this maximum level regardless of their “true” ventilation requirements. To ensure that production requirements are always fulfilled even when mines undergo significant changes (i.e. new ore mineralization opens), they usually maintain more “available” production stopes than it would be required by the planned production rates. The mines provide secondary ventilation to all these production stopes even to those which for various operating reasons become temporarily inactive. Furthermore, the auxiliary delivery systems within the production and development workings are also designed for maximum ventilation requirements, namely for the maximum length of the ducting system plus additional air volumes to compensate for potential leakage from the auxiliary ducts. As a result, during the early stages of the mines’ operating life, the total intake air volumes supplied by the primary and auxiliary ventilation systems could be well in excess of their “true” air volume requirements.  1.3  Thesis Objectives  The growing demands on ventilations systems and their importance on overall mine economics require a different approach to system design and management. With this in mind this thesis will focus on the follow objectives: •  Develop a new methodology to measure the efficiency of underground ventilation systems of large and highly mechanized metal mines.  •  Develop a new ventilation system design criteria based upon a mine’s “life-cycle” airflow demand schedule determined by means of process simulation.  •  Determine the economic and environmental benefits of this novel ventilation design criteria versus traditional ventilation design by integrating mining process simulation and ventilation simulation. 4  •  Demonstrate the efficacy of this design approach by evaluating the benefits of activity based ventilation management (ventilation-on-demand) versus traditional ventilation management.  1.4  Limitations of this Research  In order to selectively deliver appropriate air volumes when, where and only as long as mining operations are performed, an underground operation would need to develop and install an “Automated Ventilation Control System”. This thesis will not focus on the development and installation of automated ventilation control systems due to the fact that the employment of such control system largely depends upon particular characteristics of an underground operation and the level and type of its already installed communication infrastructure (e.g. hard-wired, wireless).  1.5  Organization of this Thesis  Chapter 1 discusses the current mining trends with respect to the Canadian underground metal mines such as increasing depth, mechanization, environmental concerns and the effects these trends will generate on the ventilation system. In this chapter, current ventilation design deficiencies are highlighted and the objectives of the thesis are presented and discussed. Chapter 2 summarizes past and current methods of solving and balancing mine ventilation networks and presents an overview of current communication infrastructures and event-based ventilation control systems. Chapter 3 describes the procedures that were followed in order to accomplish the objectives of this research work. Chapter 4 includes a brief description and characterization of the mining methods employed by Canadian mines, information which was used to identify which type of mines would benefit the most from this new ventilation design concept. Chapter 5 introduces a new ventilation system evaluation method, in which the magnitude of ventilation redundancy is used to measure the efficiency of mine ventilation systems. Chapter 6 provides background information with respect to the potential benefits of utilizing discrete-event mining process simulation as a decision making tool within the mine design process. This chapter introduces AutoModTM its building modules and process simulation capabilities. Chapter 7 describes the development of the “Mining Process” model of a multi-level mining block. Chapter 8 discusses the underperformance of the current methods of estimating airflow requirements in large and deep metal mines and their potential 5  impact on the business operation and the underground working environment. In this chapter the “life-cycle” airflow demand schedule of a highly mechanized metal mine is determined based upon output data generated through mining process simulations. Chapter 9 presents the subsequent steps of solving and balancing the primary and auxiliary ventilation systems of a deep metal mine (Mine A), for traditional and activity based airflow requirements by means of ventilation simulation. Furthermore, the mine’s fan power requirement and ventilation operating costs for traditional versus activity based airflow requirements are compared and discussed. Chapter 10 provides a cost estimate of an event-based control system, which includes the communication infrastructure, vehicle tracking and tagging system, a web-based energy management system and compliance monitors. Presents and discusses the economic and environmental benefits of an activity based ventilation system versus traditional ventilation system.  6  2.  LITERATURE REVIEW  2.1  Underground Ventilation Systems – Background Information  The design of an underground ventilation system is a complex process due to many interacting development, production and mine support systems. It is important that ventilation system design and other mine support infrastructures must not be treated in isolation during the overall mine design process. However, in the past it has often been the case that the sizes of mining equipment along with the required rate of mineral production and ground stability considerations have dictated the layout of a mine without, initially, taking into account ventilation demands. This usually results in inefficient ventilation systems with high capital and operating costs. In some cases, the results of inadequate planning and ventilation system design deficiencies can result in production losses, high costs of reconstruction, poor environmental conditions and tragic consequences to the health and safety of the workforce [McPherson, 1993]. Generally, almost every underground operation is unique in its geometry, extent, geological surroundings and environmental pollutants. Therefore, the corresponding patterns of airflow through the airways of these underground operations are quite variable. However, there are certain features and essential characteristics that are common, which permit classification of underground ventilation systems and subsystems. In mines, the “primary” ventilation system includes all vertical, near-vertical and horizontal airways through which the fresh air is delivered along the main airways of mine as well as the airways through which the contaminated air coming from the working areas is exhausted to the surface. For example, in metal mines, fresh air is usually delivered down intake shafts, fresh air raises (FARs) and the ramp system to the main levels, sublevels and transfer drifts. Airflow across each level is usually controlled by doors, regulators or booster fans. From here, the fresh air is picked up and delivered to the production draw-points, drilling areas and development faces through auxiliary ducts with the assistance of the auxiliary fans. From the working areas the contaminated air is then directed along the return airways to the mine’s return system, which usually consist of an exhaust shaft and return air raises (RARs). Ideally, an “auxiliary” ventilation system which can consist of a large number of auxiliary ducting systems (steel/fibre/fabric) and auxiliary fans through which fresh air is delivered to the production areas should have no impact on the distribution of airflow within the mine’s 7  primary ventilation system. This is important because it would allow the auxiliary ventilation system to be planned and designed independently from the mine’s primary ventilation system. Auxiliary ventilation systems that are used to supply fresh air to the working faces and dead-end headings can be classified into three basic types, namely: “line brattices”, “fan & duct” and “ductless” systems. Generally, the use of “line brattices” is common in underground operations employing the room-and-pillar mining method and its variations. Two major disadvantages of employing a “line brattices” system are the relatively high frictional and shock pressure losses at their “inbye” locations and potential for increased leakage throughout the auxiliary system [Wallace, 2001]. The most common type of auxiliary ventilation system employed in Canadian metal mines is the “fan-duct” system, in either forcing or exhausting mode. In metal mines, a “forcing” fanduct auxiliary ventilation system can quickly and effectively deliver fresh air to the face of the development and production workings in order to dilute and remove the pollutants generated by the mining operations. Furthermore, as this system is under positive pressure, from the fresh air pickup location to the face of the production workings then a less expensive auxiliary duct (i.e. fabric ducts) can be employed. The main disadvantage of a forcing fan-duct auxiliary setup is that the mining activity related pollutants that are added to the air at the face will affect the full length of the dead-end heading and the return airways as the contaminated air passes back along these airways at relatively low velocity rates [Hartman et al, 2006]. Where dust is the main contaminant, an “exhausting” fan-duct auxiliary ventilation setup may be preferred. With this setup, the polluted air at the face is immediately drawn into the auxiliary ducting system at the face, allowing fresh air to flow along the dead-end heading or along the production draw-point’s haulage drift. However, an exhausting fan-duct auxiliary system would need to employ more expensive rigid ducts (i.e. steel ducts) or spiral type flexible ducts [Hartman et al, 2006]. To combine the advantages of both forcing and exhausting fan-duct auxiliary systems, an “overlapping” fan-duct arrangement can be used. This type of auxiliary system is usually used in mining operations where continuous miners or tunnelling machines are employed. An important advantage is that in these mines an overlapping fan-duct auxiliary system can be mounted directly on the mining equipment. Consequently, auxiliary ventilation installation and maintenance costs can be significantly reduced. Details regarding auxiliary ventilation 8  systems and schematics of forcing, exhausting and overlapping fan-duct delivery systems are presented in the “Subsurface Ventilation and Environmental Engineering” [McPherson, 1993] and “Mine Ventilation and Air Conditioning” [Hartman, 1991] textbooks. The applicability of “ductless” auxiliary ventilation systems were tested and evaluated by CANMET-Mining and Mineral Sciences laboratories (CANMET-MMSL) in a potential “manless” mine where the production and development operations such as drilling, blasting, ore/rock haulage were performed using mining equipment remotely operated from a control room located on surface. With no workers underground, who under traditional operating conditions maintain these fan-duct auxiliary infrastructures, the fresh air to the working areas needs to be delivered without the use of auxiliary ducts. One such identified method was the use of auxiliary jet-fans. Various jet-fan installation arrangements were tested by CANMETMMSL using tracer gas. The results were used to evaluate and maximize the airflow penetration depth created by the jet-fans [Baiden, 2001], [Kocsis et al, 2001].  2.2  Review and Quantification of the Current Ventilation Design Criteria in Canadian Mines  In Canada, provinces and territories are responsible for occupational health and safety in mines, with the exception of Crown Corporations and uranium mines that fall under the federal jurisdiction. The preparation of Canadian standards for application to diesel equipment in mines began in the late 1970’s and was an initiative of the provincial inspectorates in forming the Canadian Safety Association (CSA) Standard Committee with representatives from all parts of the industry. Two standards resulted from this work, published by the CSA, one for non-gassy underground mines (CSA, 1990), and the other for gassy underground coal mines (CSA, 1988). Both standards describe the technical requirements and procedures necessary for design, performance and testing of new or unused diesel-powered machines in underground mines. The main feature of the CSA Standard is the application of a quality criterion by which to assess the comprehensive toxicity of diesel emissions in order to effectively prescribe ventilation requirements for a certified machine. The use of a comprehensive criterion is important because of the substantial changes in the individual concentrations of toxic constituents produced by different engines. Table 1 summarizes ventilation requirements for diesel engines used in Canadian underground metal mines for provinces and territories which 9  adapted diesel engine certification standards (CSA) and/or based upon mine safety and health (MSHA) regulations [Gangal & Grenier, 2002]. Table 1: Ventilation Requirements in Canadian Mines Based on Diesel Engine Certification [Gangal & Grenier, 2002]  Province/Territory British Columbia Alberta Saskatchewan Manitoba Ontario Quebec New Brunswick Nova Scotia Newfoundland & Labrador Northwest Territories & Nunavut Yukon Territories  Certification Standards NOTES CSA MSHA Ventilation as per CSA standards. Yes Ventilation according to CSA standards. Yes Minimum ventilation 0.032 m3/s/kW in active headings Yes Minimum ventilation 0.063 m3/s/kW of diesel power Ventilation as per CANMET or MSHA standards For non-approved engines ventilation 0.092 m3/s/kW. Yes Yes Minimum ventilation 0.045 m3/s/kW of diesel power Minimum ventilation 0.063 m3/s/kW. Ventilation as per CANMET or Part 31/32 of MSHA For non-approved engines ventilation 0.092 m3/s/kW Yes Yes Minimum ventilation 0.045 m3/s/kW Certification for engines above 75 kW is required Yes Yes Minimum ventilation 0.067 m3/s/kW of diesel power Or a certificate that engine meets various conditions Yes Yes Minimum ventilation 0.060 m3/s/kW Ventilation as per CANMET or MSHA standards Yes Yes Minimum ventilation 0.047 m3/s/kW of diesel power Minimum ventilation 0.060 m3/s/kW of diesel power -  -  Requires approval from testing laboratory Minimum ventilation 0.051 m3/s/kW of diesel power  2.2.1 Heat Sources in Underground Mines Heat is transferred to the airflow from a variety of sources. Within the soft-rock mines (potash, gypsum, salt, coal), the airstream itself is sufficient to remove the heat that is produced during the production processes. However, in deep Canadian metal mines, where heat generation can become the dominant environmental problem, refrigeration may also be required in order to maintain adequate underground working conditions. The four major heat sources in Canadian mines are: •  Auto-compression  •  Mining equipement  •  Strata (geothermal gradient)  •  Production/development blasting 10  Auto-compression Air entering the mine through a vertical airway is compressed and heated as it flows downward.  Auto-compression occurs when potential energy is converted into thermal  energy. If there is no interchange in heat or moisture content of the air in the shaft, the compression occurs adiabatically, with the temperature rise following the adiabatic law [Howard, 1982]. Precise calculation of ΔT due to auto-compression can be challenging, mainly because of the non-adiabatic airflow that usually occurs in mine shafts and vertical openings. This is because in addition to the heat generated by auto-compression, pickup of both strata heat and moisture along an intake shaft can occur concurrently. Auto-compression may also be masked by the presence of other heating or cooling sources located in or near the shaft, such as air and water lines. Mining Equipment Vehicles operating throughout the mines, electrical transformers and fans are all devices that convert input power, via useful effects (i.e. work), into heat. Nearly all the energy consumed by a piece of mining equipment adds heat to the airflow, since the power losses and most of the work performed are converted directly or indirectly through friction into heat. In Canadian mines, increased mechanization, equipment utilization and increasing power demands have resulted in such underground equipment being considered to be one of the major sources of heat. For electrical equipment, the total heat produced is equivalent to the rate at which power is supplied. In comparison, the internal combustion engines of diesel equipment have an overall efficiency of only one-third of that achieved by electrical units. Hence, diesels will produce nearly three times more heat than electrical equipment of the same mechanical work output. Further to this, approximately one-third of the heat generated by diesels appear as heat from the radiator and machine body, one-third as heat in the exhaust gases and the remaining one-third as useful shaft power, which is also converted into heat through frictional processes. A major difference between diesel and electrical equipment is that diesel equipment produces part of their heat output in the form of latent heat [Partyka et al, 1998]. Equipment testing performed in underground metal mines has shown that the litres of water emitted per litre of fuel consumed may vary from 1 to 10, depending upon the diesel engine’s cooling system and emission control devices [Sarin et al, 1997]. 11  Strata (geothermal gradient) When air passes through an airway, its temperature usually increases. This is caused by natural geothermal heat being conducted through the rock towards the airway, then passing through the boundary layers that exist in the air close to the rock surface. This heat transmission is greater in working areas where rock surfaces were freshly exposed (e.g. production blast), where the rock surfaces are often warmer than the air. However, these rock surfaces will cool in time to an equilibrium point where the temperature of the rock surface is only a fraction of a degree centigrade higher that the temperature of the air [MVS, 1997]. The amount of heat transferred to the air from the strata is also a function of depth. The increase in strata temperature with respect to depth is known as the “geothermal gradient”. In practical utilization, it is often inverted to give integer values and is referred to as the “geothermal step”. As one progresses downwards through a succession of geological formations, the geothermal step will vary according to the thermal conductivity and diffusivity of the local material. The age of the rock, its thermal properties, and its proximity to recent igneous activity can all affect the geothermal gradient. Even within a single mining district, it can vary and seldom be considered a constant value [MVS, 1997]. Explosives As over 75% of the energy released by a blast is liberated in the form of heat, blasting can also be considered a significant source of heat. Part of the heat generated by development and production blasts is stored within the blasting fumes, which can cause a peak heat load onto the ventilating air. The remainder is stored within the fragmented ore. The amount of heat stored within the fragmented ore can depend upon the extraction method and in particular upon the quality of ore fragmentation. With the use of electronic detonators and computercontrolled blasting, as much as 40-50% of the heat produced by production and development blasts could be rapidly removed along with the blasting fumes [Baiden, 2001].  2.2.2 Health and Safety Concerns in Canadian Mines - Exposure Limits and Guidelines It is a recognised fact that diesel powered mining equipment has allowed the mining industry to achieve tremendous improvements in productivity over the past 30 to 50 years. This is because the diesel engine is an economical source of power and the diesel powered machines 12  are rugged, mobile and versatile. On the negative side, the toxicity of some diesel exhaust gases is well documented. In addition, increasing evidence is becoming available about the potential carcinogenicity of the solid fraction of diesel exhaust known as “diesel particulate matter” or DPM [Grenier et al, 2001]. In 1988, the National Institute for Occupational Safety and Health (NIOSH) issued a report which stated that diesel exhaust as a whole is a suspected occupational carcinogen [NIOSH, 1988]. In 2000, the U.S. Environmental Protection Agency (EPA) classified diesel exhaust as “likely to be carcinogenic” [EPA, 2000]. In 1990, the Canadian ad-hoc Diesel Committee published a guideline suggesting that exposure to DPM should be measured using the “respirable combustible dust” (RCD) method and be limited to 1.5 mg/m3 over a normal 8hour shift [CDC, 1990]. This exposure value has since been adopted by most mining provinces in Canada. The American Conference of Governmental Industrial Hygienists (ACGIH) is the organization that publishes the Threshold Limit Value (TLV) guidelines on an annual basis. The ACGIH guidelines are not regulated limits by themselves but are used in many parts of the world as legal exposure limits by regulators. In 1996, the ACGIH published a notice of intended change in which a DPM-TLV of 0.15 mg/m3 was suggested [ACGIH, 1996]. In 1998 the ACGIH further reduced this proposed TLV to 0.05 mg/m3 [ACHIH, 1998]. Both of these exposure limits were based on the measurement of total carbon similar to the MSHA rule. In 2001, the ACGIH suggested a limit of 0.02 mg/m3 based on the measurement of “elemental carbon” only [Meets et al, 2000]. This proposed exposure limit of 0.02 mg/m3 of elemental carbon was later dropped. These guidelines and exposure limits are summarized in Table 2. Table 2, shows that the TLV limits adopted by the mine safety and health committees and regulating agencies since 1990, have been significantly reduced and there is no doubt that the existing limits of exposure will be reviewed. In order to comply with the newly suggested exposure limits the airflow requirements for production workings will significantly rise. An immediate consequence of this would be a dramatic increase in ventilation energy consumption and operating costs. This again, shows the importance of managing the air volume that is delivered underground.  13  Table 2: Proposed/Adopted Exposure Limits and Guidelines within the Mining Industry [Gangal & Grenier, 2002] Agency/Committee/Regulator  Date Submitted  Exposure Guideline/Limit  Substance Measured  Canadian ad hoc Diesel Committee (presently in effect in most Canadian mining provinces)  1990  1.50 mg/m3  RCD  Mine Safety and Health Administration  2001  0.40 – 0.16 mg/m3  TC  Switzerland, tunnelling American Conf. of Gov. Industrial Hygienists  N/A 1996  3  TC  3  TC  3  0.20 mg/m 0.15 mg/m  American Conf. of Gov. Industrial Hygienists  1998  0.05 mg/m  TC  Germany, tunnelling  N/A  0.10 mg/m3  EC  American Conf. of Gov. Industrial Hygienists*  2001  3  0.02 mg/m  EC  Mine Safety and Health Administration (MSHA)  2008  0.16 mg/m3  TC  * This proposed exposure limit was later dropped  2.3  Ventilation Network Analysis – Background Information  An important component in the design of a new mine is the determination and distribution of the required air volumes to the production areas and throughout the mine. During the life of an underground operation, it is important to plan ahead in order that new fans, intake and exhaust raises and other new infrastructures are available in a timely manner in order to efficiently provide the required airflow to the underground workings. Due to the dynamic nature of the underground operations with new orebodies and their production blocks under continuous development as older ones approach the end of their operating life, mine ventilation planning would also need to be a continuous process [McPherson, 1993]. Ventilation network analysis is a generic term for a family of techniques that enable quantitative determination of the distribution of airflow and the operating duties of the fans in a ventilation network where the resistance of all branches are known. In a given network there are, theoretically, an infinite number of combinations of airway resistances, pressure generators and regulators that will give a desired airflow distribution. However, practical considerations can limit the number of acceptable solutions. The techniques of network analysis that are applicable and useful for industrial applications must be flexible and converge to a solution in a short time period in order to allow multiple alternative solutions to be investigated [McPherson, 1993].  14  Before the mid 1950s, there was no practicable means of conducting detailed and quantitative ventilation network analysis for complete mine systems. Ventilation planning was carried out using hydraulic gradient diagrams formulated from assumed airflows or simply, based on experience and intuition of the ventilation engineer. The first viable electrical analogue computers to simulate ventilation networks were produced in the United Kingdom (Scott et al, 1953), then in the United States (McElroy, 1954). These analogues basically enabled Ohm’s law for electrical conductors to emulate the square law of ventilation networks. In the 1960s, ventilation simulation programs for mainframe digital computers started to appear. These proved to be more versatile, rapid and accurate and their employment soon dominated ventilation planning and system design. In the 1980s, the enhanced power and relatively reduced cost of personal computers led to the development of mine ventilation programs, which also incorporated the use of graphics [McPherson, 1993].  2.3.1 Methods of Solving Mine Ventilation Networks There are basically two sets of methods of solving fluid networks: •  “Analytical” methods, which involves the formulation of the governing laws into sets of equations that can be solved analytically in order to obtain a solution, and  •  “Numerical” methods, which become very popular with the availability of personal computers, which were then used to solve the equations through iterative procedures.  Analytical Methods The “equivalent resistances” of a ventilation network is the most elementary method of solving and balancing a ventilation system. If two or more airways are connected in series or in parallel then each of those sets of airway resistances may be combined into a single “equivalent” resistance. The concept of equivalent resistances is particularly useful when combining two or more airways that run parallel to each other. Through this technique, thousands of such branches that may exist in a mine can be reduced to hundreds, thus simplifying the network schematic, reducing the amount of data needed to be processed and consequently, minimizing the time of solving and balancing a ventilation system [McPherson, 1993]. Kirchhoff’s first law states that the mass flow entering a junction equals to the mass flow leaving the junction: 15  ∑M  = 0 or ∑ Qρ  j  Where:  (4)  j  M = mass flow (kg/s) positive and negative entering junction “j” Q = air volume (m3/s) ρ = air density in (kg/m3)  In underground ventilation systems, the variation in air density around any single junction can be considered negligible, therefore:  ∑Q  =0  (5)  j  Kirchhoff’s second law applied to ventilation networks is that the algebraic sum of all pressure drops around a closed mesh in the network must be zero:  ∑ (P − P ) f  Where:  - NVP = 0 (Pa)  (6)  P = frictional pressure drop (Pa) Pf = rise in total pressure across the fan (Pa) NVP = natural ventilation pressure (Pa)  Kirchhoff’s laws allow us to write down equation (5) for each independent junction of the network and equation (6) for each independent mesh. Solving these two sets of equations will result in determining the airflows (Q) in the branch that satisfy both laws. The direct application of Kirchhoff’s laws to a mine ventilation network could result in several hundred equations that are needed to be concurrently solved [McPherson, 1993]. Numerical Methods The method that is most widely used in computer programs for ventilation network analysis was originally developed for water distribution systems by Professor Hardy Cross at the University of Illinois in 1936. This was modified and further developed for mine ventilation systems by D. R. Scott and F. B. Hinsley at the University of Nottingham in 1951 [McPherson, 1993]. The Hardy Cross method is a method of successive approximation in which the airflows in a ventilation network are initially estimated then adjusted towards their true values iteratively through a series of corrections.  16  2.3.2 Ventilation Network Simulation Programs The primary purpose of a mine ventilation simulator is to predict the airflow distribution and pressure differentials throughout a network and produce numerical results that approximate those that would be given by a real system. There are three major factors that govern the accuracy of a mine ventilation program: •  The accuracy with which the processes are represented by its corresponding equations (e.g. P = RQ2),  •  The accuracy of the data that was used to define the ventilation model (e.g. airway resistances), and  •  The accuracy of the numerical procedure to converge to a solution  The parameters and design elements that are needed to develop the ventilation model of an underground operation are: •  Numerical data used to define each airway (branch) of the ventilation network  •  A schematic representation of the ventilation network, which consists of junctions and delineating branches in a closed circuit. Each branch can represent a single airway, leakage path, or a group of airways combined into an equivalent branch.  •  Data needed to define the locations and characteristics of the mechanical ventilation devices (i.e. fans, doors, regulators, bulkheads).  After the “model development” phase, the current airflow requirements for the main production levels (i.e. drilling and haulage levels) and sublevels are evaluated. The basic stipulation in determining the airflow requirements in mines where diesels are used is that there should be sufficient airflow to dilute exhaust gases and particulates below their TLV values. In metal mines, the criterion that is most widely used to estimate the airflow needed to dilute the diesel exhaust gases is based on the rated output power of the diesel equipment, involved in the mining process. The ventilation system is then solved and balanced by means of network simulation techniques [Kocsis et al, 2003]. Based upon current ventilation design criteria, the traditional design sequence of a mine ventilation system using a ventilation program is shown in Figure 1 [Kocsis et al, 2003].  17  Figure 1: Traditional Ventilation Design Practice [Kocsis et al, 2003] Ventilation Ventilation Survey Survey  Mining Mining Plans Plans  A A  Ventilation Ventilation Simulator Simulator STOP STOP  Ventilation Ventilation Database Database  Model Model Development Development  New New Ventilation Ventilation Scenario/Analysis Scenario/Analysis  Determine Determine Airflow Airflow Dev./Production Dev./Production  Balance/Solve Balance/Solve Ventilation Ventilation Network Network  Modify/Update Modify/Update Ventilation Ventilation Network Network  Airflow Airflow Distribution Distribution Satisfied? Satisfied?  Determine Determine New New Airflow Airflow -- Heat Heat Management Management  Climatic Climatic Conditions Conditions Appropriate? Appropriate?  Check Check U/G U/G Climatic Climatic Conditions-ClimSim Conditions-ClimSim  2.4  Ventilation Ventilation System System Optimization Optimization  Ventilation Ventilation System System Optimized? Optimized?  Print Print Solution Solution Report Report  Perform Perform Modifications Modifications New New Fans Fans Selection Selection  A A  Ventilation Control Systems  A relatively new approach used to minimize costs associated with variation in ventilation requirements during a mine’s operation life is the use of an activity based ventilation control system. To provide activity based ventilation delivery to the production workings, the mines would need to develop and install a “Ventilation Control System”. The basic flowchart of a generic ventilation control system is shown in Figure 2, and includes the following main building blocks: •  Vehicle Identification and Tracking System: To selectively deliver appropriate air volumes to the production workings according to various mining activities it is essential to know where, when and for how long these air volumes are needed. In highly mechanized metal mines, this means knowing where production, development and service vehicles are along with their characteristics. This could be achieved through the use of a vehicle identification and tracking system. These systems are designed such that ventilation and equipment parameters can be transmitted through 18  dedicated hard-wired or wireless communication systems, or a combination of both [Meyer, M.A., 2008]. •  Ventilation Controls: Within the primary and auxiliary ventilation systems there are two types of controls, active controls (i.e. fans) and passive controls (i.e. doors & regulators). Presently, both controls can be installed easily and cost effectively. Primary and auxiliary fans can be fitted with soft starters and variable frequency drives (VFDs) and their airflow delivery controlled through either variable pitch or rotational speed. Furthermore, most commercial ventilation doors can be adapted in order to allow variable amounts of airflow along the development and haulage drifts. Some ventilation doors can also act as regulators when partially open; alternatively, specific regulators with a finer control of volume may be required [Fink & Beatty, 1987], [Allen & Keen, 2008].  •  Compliance Monitors: To ensure that an activity based ventilation system provides and maintains adequate working conditions within the production workings, two types of environmental monitoring systems may be required. Firstly, the required air volume at the face of the developments, drilling locations and stope draw-points must be guaranteed. For this, non-intrusive ultrasonic airflow sensors are best suited along the main haulage drifts. Secondly, the quality of the air delivered to the production workings must also be ensured. For this, a variety of single and multiple gas monitors are presently available with the ability to measure the by-products of diesel activity and blasting pollutants. Continuous DPM monitors are now available. These monitoring units can be installed within the production draw-pints, ore/waste pass areas and along busy haulage drifts. To ensure the equipment operators are always exposed to safe underground climatic conditions, the continuous DPM monitors can also be installed on the mobile production equipment.  •  Data Management System: This includes the process logic of the ventilation control system by which information is processed and outcome determined. Presently, there are several data management systems available on the market; among them Bestech’s “Energy Management System” (NRG-1) and Simsmart’s “Optimized Mine Ventilation On-Demand” (OM-VOD) can be mentioned.  19  •  Communication System: Information gathered from the vehicle identification and tracking system and compliance monitors is transmitted to the data management system, and from here to the main and local PLCs through the mine’s communication system. A mine-wide communication infrastructure can be based upon a combination of “Local Area Networks” (LAN), Ethernet and “Allen-Bradley Data Highway Plus” (DH+) networks [Meyer, 2008].  Figure 2: Generic Ventilation Control System Components [Kocsis et al, 2004] Selective Flow Distribution  Ventilation Controls  Data Management System  Vehicle Identification and Tracking System  Compliance Monitors  Working Areas  Within the underground operations, three levels of ventilation control systems can be developed and installed, namely: •  Time-of-Day control system  •  Event-Based control system, and  •  Real-Time control system (also termed as Dynamic VOD System)  With a “Time-of-Day” control level, the main and local PLCs that are controlling the fans within the primary and auxiliary ventilation systems are programmed to be turned “Off” and the airflow regulators adjusted accordingly at predetermined times such as at the end of the production shifts or during weekends. However, the gas monitors would have the ability to override the pre-programmed schedules in case one of the monitored gasses (e.g. CO, CO2, NO, NO2) exceeds its prescribed threshold limit values (TLVs) [Hardcastle et al, 2008] [Meyer, 2008]. With an “Event-Based” control level, the fans and airflow regulators are controlled by the Data Management System (DMS). Information to the DMS is provided by the vehicle identification and tracking system which is able to identify the type and location of the mining equipment. This can be accomplished using identification “tags” carried by both 20  mining equipment and personnel. The tags are coded with data such as type of equipment (diesel/electric), power (hp), required air volume per rated power (m3/s per kW) and equipment/personnel identification number. The information is processed into air volume requirements by a VOD software package operating on the DMS server. These data are transmitted to the surface control room and the main PLCs located on the production levels. From the main PLCs the controlling instructions are transmitted to the local PLCs which in turn will adjust the operating speed of the auxiliary fans each equipped with variable frequency drives (VFDs) until the required air volumes are satisfied. Air volume information from airflow sensors installed inside the auxiliary ducting systems is transmitted back to the DMS to ensure that adequate air volumes are delivered to the production workings. Again, in case that one of the monitored gasses exceeds its TLV, the local PLC gathering air quality information in a production area would have the ability to override the instructions transmitted by the DMS and instruct the auxiliary fans associated to that particular production area to operate at their maximum capacity until the quality of the air is restored [Nenesen & Lundkvist, 2005] [O’Connor, 2008]. A “Real-Time” control level would only be accomplished with the installation of a fully automated ventilation system where information transmitted from the tag readers, environmental monitors and airflow sensors are transmitted to the DMS. This is processed in real-time by VOD software and a ventilation program running in parallel on the DMS server. As opposed to air volume instructions carried by equipment and personnel tags, with this level of control the air volume requirements for the production workings and throughout the mine are determined by a ventilation program. With this approach, any changes within the mines primary and auxiliary ventilation systems such as an additional pressure requirement from a particular fan(s) due to a temporary restriction along an airway is processed by the ventilation program in real-time and the information is transmitted to the main and local PLCs [Tonnos & Allen, 2008].  2.4.1 Examples of Event-Based Ventilation Control Systems Various levels of underground ventilation control systems are presently being installed and employed world-wide. Within the mines, the level of these control systems usually depends upon the complexity of their primary and auxiliary ventilation systems, number of primary  21  and auxiliary fans, employed mining method, level of mechanization and existing communication infrastructure. In North America among the mines that are presently installing communication infrastructures able to provide an event-based level of control, two mines can be considered and their control systems discussed, as follows: •  Ventilation on Demand System at Vale Inco’s Creighton Mine, Sudbury, Canada  •  Equipment Tracking and Ventilation Control System at Barrick’s Goldstrike Mine, Nevada, USA  In Europe and Australia, among the mines that are already delivering demand based air volumes to the production workings and have implemented real-time airflow monitoring systems, the following two projects are reviewed: •  Ventilation on Demand System at LKAB’s Malmberget Iron Ore Mine, Sweden  •  Real Time Airflow Monitoring and Control System at Capcoal Central Colliery and Cannington Mine, Queensland, Australia  Example 1: Event-Based Ventilation on Demand System at Vale Inco’s Creighton Mine, Sudbury, Canada At the Creighton mine, energy conservation initiatives such as a time-of-day ventilation control system have already been implemented. Further to this, a feasibility level study [Hardcastle, et al] which looked at the mine’s ventilation system showed that further reductions in energy consumption can be achieved with the development of an event-based control system. Consequently, the mine initiated a phased approach to the development of this control system. The first phase included a pilot project with the main objective to evaluate a practical methodology for the development of an event-based ventilation control system on the 2340m level. This project involves the installation of variable frequency drives (VFDs) on the auxiliary fans in order to allow variable air volumes to be delivered to the production workings by varying the operating speed of the auxiliary fans. The schematic of this eventbased control system is presented in Figure 3. The development and implementation of an activity based control system requires the following components [O’Connor, 2008]: 22  •  Dedicated server hardware and VOD software  •  Vehicle identification and tagging system (tags, readers, exciters)  •  Airflow sensors and air quality monitors  •  Programmable logic controllers (PLCs) & human machine interface (HMI)  Figure 3: Creighton mine’s ventilation on demand schematic [O’Connor, 2008]  The required air volumes for the production workings are determined by the VOD software operating on the process control server. The tags on the mining equipment carry information such as equipment identification number, power (kW) and required air volumes based upon diesel exhaust dilution requirements, information which is transmitted to a tag reader as soon as the mining equipment arrives within a tag reader’s zone of influence. This information is processed by the VOD software into air volume requirements and transmitted to the main PLC which controls the auxiliary fans and associated regulators. Should the communication system between the process control server and the main PLC fail, then the local PLCs  23  gathering data from the airflow sensors and air quality monitoring units would instruct the auxiliary fans according to the mine’s predefined “fail-safe” protocol [O’Connor, 2008]. Presently, this event-based ventilation control system on Creighton Mine’s 2340 level is being evaluated. The development of this ventilation control system involved the collaboration of several disciplines including ventilation engineers, electrical engineers, programmers, operating personnel and equipment suppliers. The success of this pilot project can lead to the next phase which will involve the installation of an event-based ventilation control system mine-wide. Example 2: Equipment Tracking and Ventilation Control System at Barrick’s Goldstrike Mine, Nevada, USA The underground division of Barrick’s Goldstrike mine is made up of several mining zones including Meikle, Rodeo, Griffin and others. There are two major haulage drifts connecting Meikle and Rodeo through the Griffin zone. Ramp systems connect all mining levels. The mine has two ventilation intake shafts, one at Meikle and one at Rodeo Mine and three exhaust shafts, two at Meikle and one at Rodeo Mine. A decline haulage drift from the BetzePost pit enters the Rodeo Mine from the south as another fresh air intake. The lowest mining level is currently 610 m below surface at an elevation of 1097 m above sea level [Meyer et al, 2008]. The amount of air exhausted out of the mining zones is 1,080 m3/s. Meikle Mine has two 2.7m diameter 1,300 kW centrifugal exhaust fans on the main exhaust shaft and one 1.9m diameter 520 kW centrifugal fan on the top of ventilation raise, which together pull 517 m3/s. Rodeo Mine located at approximately 1.6 km to the south has two 2.7m diameter 1,120 kW variable speed axi-vane fans located at the top of the exhaust shaft pulling 564 m3/s. In addition to the surface there are numerous axi-vane booster fans located underground ranging from 75 to 375 kW. These booster fans are essential in distributing airflow to the various levels and production areas of the mine [Meyer et al, 2008]. Existing Communication System Since 1995, an extensive fibre optic network has been installed from surface and throughout the mine to all main production levels. Several copper or twisted pair networks are also installed with media converters and routers to interface the local devices with the fibre optic 24  network. The mine’s main communication network is based upon a combination of Ethernet, Allen-Bradley Data Highway Plus (DH+) and “DeviceNet” which interfaces to numerous PLCs throughout the mine. Fixed plants such as the shotcrete plant, backfill plant, loading pockets, crushers, mine hoist and refrigeration plant are all connected to the mine’s main communication network through local PLCs. All booster and auxiliary fans are powered from motor control centres (MCCs) or from mobile load centres (MLCs), and can be remotely controlled [Meyer, M.A., et al, 2008]. The flowchart of Barrick’s underground division equipment tracking and ventilation control system is shown in Figure 4. Figure 4: Barrick’s Equipment Tracking and Ventilation Control System  Ventilation Monitoring System Numerous gas monitors have been installed at the mine in recent years. The primary purpose of these monitors is to detect high levels of CO. In other areas of the mine additional gas sensors were installed to monitor pollutants such as CO2 (from ground water), SO2 (from sulphide ore), NO and NH3 (from blasting and refrigeration). Ultrasonic type airflow monitors have been installed on several levels, ramps and main intersections. Where airflow can reverse direction, the airflow monitors were configured to provide bi-directional flow readings. Several remotely adjustable airflow regulators were installed at the bottom of the 25  main fresh air raises. The regulator actuators were connected to the mine’s “LonWorks” communication network to allow remote control of the actuators [Meyer et al, 2008]. Emergency Notification System In 2005, Barrick Goldstike Underground Division started looking at tracking vehicles to monitor location and payload information. As a result of the mining accidents in the USA in 2006 and 2007, the need to track mining personnel throughout the mines became a priority. Firstly, an emergency notification system was required to send a message to the miners in the event of a mine emergency. Secondly, personnel identification and tracking was required to track and locate miners throughout the mines. Mine Site Technologies (MST) was selected to develop and install an emergency notification system [Meyer et al, 2008]. MST’s PED emergency notification system is a one-way, through-the-earth communication system capable of sending a text message to the wearer of a PED receiver. The PED head end panel was installed at Meikle Mine’s administration building. The antenna is a continuous loop that was installed partially on surface and partially underground from Meikle to Rodeo Mine, then down Rodeo intake shaft, through Griffin and up through Meikle production shaft. MST’s Integrated Communication Cap Lamp (ICCL) was selected to replace the existing cap lamps. The ICCL houses the PED receiver and the 802.11 Wi-Fi tag. In order to track the Wi-Fi tags in the mine and to allow wireless access to the mines’ local area networks (LANs), Wireless Access Points (WAP) were installed at major intersections, plant locations, refuge stations and shaft stations. The WAPs are connected together using a hybrid fibre optic/copper cable with two or four multimode (MM) fibres for signal and two copper conductors for the 24V power. The WAP consists of a fibre switch motherboard which communicates at 100 MB/second and provides wired LAN connectivity and wireless communication. In addition a single mode (SM) fibre optic cable was also installed along the Meikle and Rodeo intake shafts, between all levels and along the main haulage drifts of both mining zones. Ethernet switches were installed on the main levels on the SM fibre optic cable, which created 1 GB/second Ethernet communication system. The tracking program selected for the emergency notification system is the “AeroScout Engine” with “MobileView” [Meyer et al, 2008].  26  Mine Dispatch System Goldstrike underground division utilizes Micromine’s PITRAM software to track, store and report a variety of mining operations. The mine’s dispatch system includes three main components: data acquisition (DA), reporting functions and administrative controls. Typical events that are recorded include equipment states, equipment location and equipment movements. Micromine is currently upgrading its tracking software to PITRAM version 3, which will utilize Microsoft SQL server, thus allowing multiple concurrent data acquisition. A further enhancement to the system will be a mobile application, which will allow the equipment operator to enter data via a touch screen or keyboard. This application will also interface with the engine management system to gather engine performance data. As a mobile vehicle comes in range of a WPM, engine performance data stored on the vehicle’s computer will be downloaded to the PITRAM data server via the WLAN/LAN networks. The vehicle tracking software operating on the PITRAM server will be used to identify and track the mining equipment equipped with an identification tag [Meyer, M.A., et al, 2008]. It is also expected that the engine performance data can be used by Bestech’s NRG-1 energy management system to instruct the local PLCs and adjust the fans and associated regulators to deliver appropriate air volumes to the production areas. The NRG-1 ventilation control system will be described in section 2.4.2. Ventilation on Demand System So, what still remains to be developed at Barrick’s Underground Division in order to achieve an event-based level ventilation control system? A mine-wide air quality monitoring system is already in place. Airflow and pollutant levels are measured in many areas of the mine in order to ensure that minimum air volumes and air quality standards are continually maintained. A wireless network and tracking system is being installed in the mine. All personnel and most vehicles will carry Wi-Fi tags. Production vehicles such as bolters, drills, loaders, LHDs and trucks will have on-board computers for data acquisition. The on-board computers through WAPs will interface with the engine management system to provide engine performance data. An energy management system is already operating and provides time-of-day level ventilation control. The tag information associated with a particular vehicle and its engine performance data can be used to determine the air volume requirement in a working area where this vehicle operates. Main airflow regulators can be adjusted to allow an 27  appropriate amount of air volume on the production levels. The main exhaust fans could use variable pitch blades or could be equipped with variable frequency drives which would allow their operating speed (RPM) to be continually adjusted in order to match variable air volume requirements throughout the mine [Meyer et al, 2008]. The only elements that are missing to achieve an activity-based ventilation control system at Barrick’s Goldstrike mine are the variable frequency drives for the auxiliary fans. To enable the auxiliary fans to provide variable air volumes to the production workings by varying their rotational speed (RPM), the mine would need to install variable frequency drives on its auxiliary fans. Example 3: Ventilation on Demand System at LKAB’s Malmberget Iron Ore Mine, Sweden The Swedish mining company, LKAB has three production sites and operates two underground mines located in Malmberget and Kiruna, in northern Sweden. The Malmberget mine has 20 ore zones, of which 10 are presently active. The main haulage levels in the mine are at 350, 600, 815 and 1000 meters below the surface. The main extraction method at the Malmberget mine is sublevel caving. The primary ventilation system of the mine includes 9 intake and 10 exhaust stations located underground and on surface. The auxiliary ventilation system includes a number of 130 auxiliary fans [Nenesen & Lundkvist, 2005]. Previous Ventilation System The original ventilation system at the Malmberget mine was built between 1960 and the mid 1970’s. An intake fresh air shaft, namely the Dannewitz F9B shaft was completed in 1975. Initially the ventilation system was planned and developed to ventilate the mine to a depth of only 600 meters. Over the years the ventilation system has been further extended on a number of occasions to accommodate changes as development and production workings deepened [Nenesen & Lundkvist, 2005]. The New Ventilation System As the production workings deepened insufficient air volumes were delivered to the production workings and production activity interruptions prompted the mine to upgrade its ventilation system. The construction of the new ventilation system started in 1999 and was completed in 2000 through the following phases [Nenesen & Lundkvist, 2005]: 28  •  Constriction of a 3.5m diameter raisbore connecting the 500mL main haulage drift with the 820mL. This phase included the installation of two booster fans, which increased the capacity of this intake system to 250 m3/s. This intake infrastructure was also equipped with an event based ventilation control system  •  Construction of a 4.5m diameter intake shaft at the Dannewitz site from surface to the 840mL and installation of two surface fans (2 x 1,300 kW) in parallel arrangement. The fresh air delivered by each fan down the intake shaft is 350 m3/s. The capacity of the heating system was increased from 12.3 MW to 18.6 MW.  •  Construction of 7 intake and 5 exhaust raises connecting the 840mL to the new haulage drift located on the 1000mL  Activity Based Ventilation Control System To optimize the use of fresh air delivered underground and minimize energy consumption an activity based ventilation control system was installed at the Malmberget mine. This control system is currently supporting mining operations at the Alliasen, Dannewitz, Parta, Vitafors and Western Fields orebodies [Nenesen & Lundkvist, 2005]. Each mining block at the Malmberget mine has its own intake and return airways (FAR/RAR). As a result, the communication infrastructure at this mine is relatively simple. This infrastructure consists of a cable network which connects the PLCs that control the surface fans and the level PLCs to the data management program which operates on a dedicated server. The air quality monitoring system consists of CO sensors connected to the communication system through LANs. Again, the CO sensors through their associated PLCs have the ability to override time-of-day or activity based instructions provided by the data management program and instruct the auxiliary and primary fans to operate according to the mine’s predetermined “fail-safe” protocol when CO exceeds its prescribed TLV [Nenesen & Lundkvist, 2005]. Each vehicle is equipped with a radio transmitter which also carries information such as the type of the vehicle, power (kW) and required air volume. As the mining equipment approaches the production area, the radio transmitter mounted on the vehicle will start the auxiliary fan. The auxiliary fans are equipped with variable frequency drives able to deliver variable air volumes to the production area. The primary fans within the ventilation system 29  are controlled by the data management program and their operating status (On/Off) depends upon the number of active auxiliary fans. A direct effect of this activity based control system was a 29% reduction in ventilation energy consumption and a 40% reduction in heating costs [Nenesen & Lundkvist, 2005]. Example 4: Real-Time Airflow Monitoring System at Capcoal Central Colliery and Cannington Mine, Queensland, Australia This research work describes efforts to characterize and mathematically model regulators and explains how this information is used to develop a computerized ventilation monitoring and simulation system able to provide real-time information for each particular branch of an underground ventilation network through linking of airflow and differential pressure sensors to a ventilation simulation program. Underground measurements and regulators tested at the University of Queensland Experimental Mine (UQEM) indicated that theoretical calculations that are used to predict the magnitude of airflow through mine regulators based upon measured differential pressures are inadequate. These theoretical calculations have limitations due to the fact that they are based on the prediction of fluid flow through a circular orifice, whereas the orifice of most mine regulators has a rectangular shape. Furthermore, there is air leakage through the frame of the regulator, which can further increase the difference between the actual air volume flowing through the regulator and the predicted air volume that is determined through theoretical calculations [Gillies, et al, 2005]. Real Time Airflow Monitoring System It is expected that in the near future a real-time mine ventilation model will be an integral part of an online mine-wide planning, monitoring and control software platform that will be continuously updated along with the mining plans. To meet this future demand, a real-time airflow monitoring and simulation system was developed. The system measures air volume or air pressure changes in selected branches of a ventilation network. These data is transmitted to a ventilation program, which in turn simulates real-time airflow distribution through all other branches of the ventilation network. The ventilation program that was selected to be linked to monitoring units was VentsimTM. This program was modified in order to accept real-time information generated by the ventilation monitoring sensors,  30  undertake network simulations and interpret key system data as well as operational changes [Gillies, et al, 2005]. Real Time Airflow Monitoring System Trial at Capcoal Central Colliery Capcoal’s Central Colliery is located in the Bowen Basin and it was the first modern underground longwall operation in Queensland, Australia. Central Colliery has an annual production of 2.5 million tonnes. The ventilation system at Central Colliery needed to be closely monitored through the installation of louver type regulators and real-time sensors at strategic locations, as the delivery of fresh air to the operating faces started to experience difficulties due to the mine’s high resistance. Eight underground remote monitoring stations using intrinsically safe differential pressure sensors were installed for this trial. Four additional remote stations with monitoring sensors for the surface fans were also included in the system [Gillies, et al, 2005]. During the trial, some problems were identified with respect to the air velocity sensors using vane anemometers, thermal mass cooling and vortex shedding techniques. These air velocity sensors were found to be difficult to re-calibrate to variable airflow requirements. However, during the trial it was found that the monitoring system had the ability to update the mine’s ventilation network model and rebalance the airflow distribution based upon real-time data provided by the monitoring sensors [Gillies, et al, 2005]. Real Time Airflow Monitoring System Trial at Cannington Mine Cannington mine is located in north-west Queensland, approximately 200 km south of Mount Isa. Cannington mine is the word’s largest single silver producer, representing approximately 6% of the world’s primary silver production. The mine was commissioned in 1997. Since then the production capacity was expanded from 1.5M tonnes/year to 2.2M tonnes/year [Gillies et al, 2005]. Initially, C-section regulators were installed and tested at each level of the mine. Due to difficulties in changing the opening of the regulator, it was decided to replace the C-section section regulators with roller-door type regulators. The height of these roller-door regulators could be adjusted using an automated control system [Gillies et al, 2005]. For the trial, nine underground monitoring stations were installed at various levels equipped with differential pressure sensors and ultrasonic airflow sensors. During this trial, due to 31  specific site operational constrains only five monitoring stations were connected online and transmitted real-time ventilation data to the VentsimTM program. Again, this trial showed that the real-time ventilation monitoring and simulation system had the ability to detect changes within the ventilation system and rebalance the airflow distribution based upon real-time data provided by the differential pressure and airflow sensors [Gillies et al, 2005].  2.4.2 Example of a Real-Time Ventilation Control System An example of a real-time ventilation control system is Bestech’s “Model Control System” (MCS), which incorporates ventilation modelling software, control system hardware and software, environmental monitoring hardware and software and a mine-wide communication infrastructure in order to provide appropriate air volumes to the production workings and throughout the mine. Example 5: Bestech’s Real-Time Ventilation Control System (MCS) The MCS architecture includes the following main components: •  Field Devices  •  Communication Infrastructure  •  Server Hardware  •  Control System Software  •  Ventilation Modelling Software  The basic block diagram of the Bestech’s MCS real-time ventilation control system is presented in Figure 5. Figure 5: Block diagram of model control system components [Tonnos et al, 2008] Field Devices  Communication Infrastructure  Server Hardware Control System  Web Interface  Modelling System  Field Devices The connection between the working areas and Bestech’s MCS system is achieved using airflow and air quality monitoring units. These monitoring units have the ability to measure the air volume delivered to the production area, the climatic conditions throughout the 32  production area and the level of pollutants generated by the mining operations. A control layer within the MCS architecture has an advanced algorithm which has the ability to provide auto-calibration for the airflow and air quality monitoring units based upon historical environmental data provided by the PLCs that are associated to these monitoring units. Furthermore, the MCS system includes safety control logic in response to a communication infrastructure downtime and in situations when the levels of contaminants within the production area exceed their TLVs. [Tonnos et al, 2008]. Communication Infrastructure The communication infrastructure allows data to be transmitted from the field devices such as air volume sensors, air quality monitoring units and tag readers to the server hardware. The MCS system is designed to utilize OPC standards for communication on a wide variety of systems ranging from the traditional leaky feeder systems to modern fibre optic systems or 802.11 wireless networks [Tonnos et al, 2008]. Server Hardware The architecture of the server hardware that enables real-time ventilation control includes a set of servers where the first server hosts the web interface, the second server hosts the databases and the third server hosts the OPC communication layer and the data processing software. This server architecture can minimize the impact of a potential hardware or software failure upon the performance of the MCS. A battery back-up system for continuous operation of the MCS during power outage is also provided. In case of a server hardware failure, the MCS would revert into a “fail-save” operating mode which can be defined by the mine operator. In this case the local PLCs gathering data from the field devices would instruct and control the primary and auxiliary fans according to the predefined “fail-safe” protocols [Tonnos et al, 2008]. Control System The control system selected for the MCS architecture is Bestech’s NRG-1. NRG-1 is a webbased control system capable of handling three control streams, namely time-of-day, eventbased and real-time. NRG-1 utilizes the inputs from equipment tags, ventilation controls and air quality monitoring units and determines whether the control commands have been executed as expected, whether an event is being triggered and whether the required air 33  volumes were delivered to the production workings. NRG-1 transmits the air volume requirements to the main programmable logic controllers (PLCs) located on each production level. The PLCs then control the fans and associated regulators to deliver the desired air volumes to the development and production workings [Tonnos et al, 2008]. Ventilation Modelling Software The ventilation simulation program selected for the MCS was VnetPCTM. This version of VnetPCTM has been updated with a structural query language (SQL) database and enhanced user interface. VnetPCTM provides the MCS with ventilation simulation results of various control changes. The control changes are executed by the solver of the ventilation program and the updated ventilation parameters are transmitted to the MCS system. The results of this real-time ventilation simulation provides the MCS system the capability to verify the control commands being executed to ensure that appropriate air volumes are delivered to the production workings. Air volume predictions through ventilation simulation performed using VnetPCTM and baseline measurements from the air volume sensors must be within 10% tolerance of each other. Errors larger than 10% would force the MCS into the fail-safe operating environment [Tonnos et al, 2008]. These examples show that in many underground operations world-wide there are adequate communication infrastructures and control programs to allow time-of-day, event-based and real-time ventilation control systems. The determination of the activity-based air volumes through discrete-event process simulation, primary and auxiliary ventilation system design and the economical and environmental benefits of such systems determined through ventilation simulation will be described and discussed in the following sections of the thesis.  2.5  Summary  The literature review has indicated that current practices for ventilation design are not optimal with respect to life of mine design. It has also demonstrated that several mines are moving towards “ventilation on demand” as a means to minimize costs associated with the current design approach. However, again there is no method to accurately assess the benefits of moving to ventilation on demand practice. Finally, the literature reveals that there is no robust indicator of the efficacy of a ventilation design after it is implemented. The remainder of this work will attempt to address these issues. 34  3.  METHODOLOGY  Within the last decade the ventilation-on-demand concept was discussed at various conferences and seminars as well as with the mine operators. Section 2.4.1 presents various levels of underground ventilation controls that were already installed world-wide with an attempt to achieve some level of an activity based ventilation delivery system. However, so far, the full benefit of the ventilation-on-demand concept has never been achieved. This was mainly due to the following reasons: •  Lack of appropriate ventilation controls and associated software able to deliver variable amounts of fresh air to the production workings, and more importantly  •  Lack of an appropriate “tool” capable to predict the potential economic and environmental benefits of a future activity based ventilation system.  The second reason led to the development of this new “tool” which could influence the mines to commit the necessary capital investments for design and installation of activity-based ventilation systems. Throughout this research work and in the course of discussions with the mine operators it became apparent that this new “tool” would also assist the candidate in the development of a new ventilation design criteria based upon a mine’s life-cycle airflow demand schedule determined through discrete-event process simulation. This activity based life-cycle airflow demand schedule would then become the basis for the design of a mine’s ventilation system. Through case study #1 and case study #2, this thesis introduces a new method which is based upon the magnitude of a mine’s potential “ventilation redundancy” that can be used to measure the efficiency of its ventilation system. The two case studies presented in this thesis were two large and complex research projects that besides the work performed by the candidate involved the input and assistance from other researchers and support from the mines, such as underground transportation for data collection and arrangements for underground ventilation surveys. It would have been impossible to do this work without the support of others. A description of the procedures that were followed to achieve the objectives of this thesis and a summary of the research work performed by the candidate and other contributors is presented for each sub-section. Furthermore, a detailed description of the individual  35  contribution of the candidate and the input and support work of other researchers and mine operating staff is presented in Appendix A.  3.1  Underground Operations Survey  This survey was necessary to identify the Canadian underground operations that would benefit the most if their ventilation systems would be designed and operated according to this new ventilation concept, where the mines’ life-cycle airflow demand is determined through mining process simulations. Within the various stages of the overall mine design process it has been observed that an important factor which can dictate the size and shape of a mine’s main openings as well as the design of its primary and auxiliary ventilation infrastructure is the mining method that will be employed. This is because the selected extraction method can ultimately dictate the layout of the production stopes, the configuration of the stope’s drawpoints, the ground support method, ore/waste transportation system and the number and size of development and production equipment. To classify the Canadian mines based upon the employed mining method and similar operating characteristics, initially, all mining methods employed by the Canadian mines were reviewed and evaluated. For each mining method, the primary and auxiliary ventilation characteristics and specific ventilation requirements (i.e. the applicability of “through-flow” versus “series” auxiliary ventilation setups) were evaluated and analyzed. The classification was based upon an extensive survey published by the Canadian Mining Journal in its “Mining Sourcebook” for years 2005, 2006, 2007 and 2008. For some mines located within the Sudbury basin (e.g. Creighton Mine, Coleman/McCreedy East Mine) and within the province of Quebec (e.g. Niobec Mine, Mouska Mine) the configuration of the production draw-points, the auxiliary ventilation setup and the applicability of this new ventilation design concept was analyzed and discussed with the mines’ engineering/ventilation department through scheduled underground visits.  3.2  Ventilation Redundancy and its Importance in Measuring the Efficiency of Underground Ventilation Systems  The main objective of a mine ventilation optimization study is to determine and improve the efficiency of the ventilation system. Unfortunately, there are few available methods that have the ability to accurately determine the efficiency of a mine’s ventilation system. Very often, 36  the “volumetric efficiency” of a mine, which is defined as the “airflow usefully employed” divided by the “total airflow delivered by the main fans” is used to gauge the efficiency of the ventilation system. The “airflow usefully employed” term is defined as the sum of the air volumes reaching the development and production faces and underground service facilities. One recent paper highlights the fact that the “volumetric efficiency” is an inadequate and poor method in determining the efficiency of a ventilations system [Chalmers, 2008]. This is because an underground ventilation system may have a high volumetric efficiency (e.g. 90%), but at the same time it can be highly inefficient in terms of frictional and shock pressure losses. Furthermore, in the case of large and heavily mechanized metal mines, due to the dynamic nature of production operations and insufficient information with respect to where and for how long the mining equipment is active, it is problematic to quantify how much of a mine’s fresh air delivered by the main fans has actually reached the mining equipment. This thesis introduces a new method where the magnitude of “ventilation redundancy” is used to measure the efficiency of a mine ventilation system. The “ventilation redundancy” term can be defined as the “total airflow delivered by the main fans” minus the “sum of the activity based air volumes” required by the production workings and service facilities. However, accurate determination of ventilation redundancy by means of traditional methods can be problematic, labour intensive and extremely time consuming. This will be highlighted in two case studies which explore the possibility of reducing the energy consumption in large and deep underground metal mines. •  Case Study #1: Determination of Ventilation Redundancy at the Creighton Mine Vale Inco  The main objective of this study was to determine the duration over which the production, development, dumping and backfilling working sites were active during the 10-hour back-toback shifts through a combination of temperature and gas monitoring. Throughout the production shifts, attention was also focused at identifying inactive time-intervals during which the mine’s intake air volume could be reduced. The findings were then used in association with ventilation simulation to determine the potential energy savings to be gained with the introduction of a “time-of-day” or an “event-based” ventilation control system.  37  The project leader appointed for this study was Stephen Hardcastle from CANMET-MMSL. Installation of temperature and humidity sensors and climatic data collection was performed by Charles Kocsis. Climatic data processing and interpretation was performed by both Charles Kocsis and Stephen Hardcastle. Ventilation modelling and determination of the potential operating cost savings for both time-of-day and activity-based ventilation controls was performed by Charles Kocsis through ventilation simulation. The reporting task was divided in two sections. The report describing the mine activity monitoring and climatic data interpretation was carried out by Stephen Hardcastle [Hardcastle & Kocsis, 2004, Report No. 04-023-1(CR)]. The report describing the ventilation modelling work and the determination of potential ventilation cost savings for “time-of-day” and “activity-based” level controls was carried out by Charles Kocsis [Kocsis & Hardcastle, 2004, Report No. 04-023-2(CR)]. Mine Activity Monitoring through Temperature Studies During 2001 and 2002, four mining activities were monitored with temperature and humidity sensors for a total duration of 125 days, namely: two draw-points of a mining block that employs the vertical retreat mining method (VRM) for 28 and 33 days, one backfilling site for 28 days, one ore-pass site for 8 days and one long-hole drilling site for 26 days. The temperature and humidity sensors were installed inside the auxiliary ducts to monitor the conditions of the intake air, within the draw-point working area and along the return airways to monitor the conditions of the return air. Details with respect to the monitoring sensors and the locations where theses sensors were installed are presented in section 5.3.2. At every monitoring location the temperature (dry-bulb) and humidity (RH) data were downloaded on a mobile computer twice a week, usually at the end of the production shifts. These climatic data was later processed into temperature and humidity trends. The temperature and humidity trends had the ability to identify the activity cycle time, the active periods, non-active periods, the start/stop times for each of these periods plus the climatic conditions generated by the mining equipment within the active areas. A sample of this monitoring program that shows the dry-bulb temperature trends at five locations around a VRM draw-point is presented in section 5.3.2. Based upon such temperature and humidity monitoring, it was possible to generate long-term (i.e. monthly) mucking activity logs.  38  Explosive Blast Monitoring Between March 26th, 2003 and April 2nd, 2003, the production blasts on the 7400 level were monitored with Draeger MultiwarnTM gas monitors fitted with a carbon dioxide (CO2) detector plus carbon monoxide (CO), nitric oxide (NO) and nitrogen dioxide (NO2) cells. On each occasion four blasting locations were monitored. These monitors were placed prior to the blast and retrieved the next day by Vale-Inco personnel. Their monitoring frequency was every 30 seconds. With respect to health and safety requirements, the longest time needed by the auxiliary system to dilute the ambient concentration of a blast gas to 50% of its 8-hours time weighted average exposure limit (TWAEV), have been used as a criterion for safe re-entry to the production area. The exposure limits used were 5,000 ppm CO2, 25 ppm CO, 25 ppm NO and 3 ppm NO2. A typical sample of a VRM blast monitoring is shown section 5.3.3. Throughout the blast clearance evaluation, CO was always the last gas to decrease below 50% of its 8-hours TWAEV limit. As a result, CO was considered the most reliable gas to be monitored in order to determine the re-entry time to the production area. •  Case Study #2: Ventilation Utilization Efficiency at Kidd Creek Mine – Xstrata PLC  In a multi-level complex base metal mine, such as Kidd Creek mine, the ventilation system may supply an air volume well in excess of what is actually required for production. This is because at any point in time during a production shift, there can be numerous levels and working areas being ventilated despite a lack of mining activity. Even in the working areas where mining operations are performed, the air volumes supplied could be greater than that required by the operating equipment. As a result, in 2005, the Kidd Creek mine’s management initiated a three stage review of the mine’s ventilation system in order to determine and minimize the ventilation redundancy within the systems in an attempt to reduce the mines energy consumption and associated costs. The first stage review was basically a brainstorming forum organized by the mine’s engineering department which involved mine personnel and various ventilation experts from Canada and abroad. The second and third stage reviews used long-term production planning  39  data and historic mining activity data, respectively, to determine the mine’s “true” intake air volume requirement and eliminate the ventilation redundancy within the system. Within the second stage, the mine’s intake air volume requirement was initially determined by Genivar Engineering, a consulting firm form the province of Quebec. Following this, the Kidd Creek mine requested CANMET-MMSL to review the work performed by Genivar engineering and the associated airflow evaluation method that was used to determine the mine’s intake air volume requirement. This review identified deficiencies with respect to the airflow evaluation method used by Genivar Engineering and acknowledged the need for a third stage review, which was carried out solely by CANMET-MMSL. The project leader appointed for this study was Stephen Hardcastle from CANMET-MMSL. Production operations schedules and mining equipment activity data was collected at the mine site by Charles Kocsis from CANMET-MMSL. Mining and equipment activity data processing was performed by Charles Kocsis and Gary Li from CANMET-MMSL. Following these, equipment activity data interpretation and reporting was carried out by Stephen Hardcastle and Charles Kocsis [Hardcastle, Kocsis & Li, 2006]. Airflow Calculation Based on Future Production Planning (2nd Stage Review) The 2nd stage review was carried out in two phases, which in this case study are referred to as “iteration #1” and “iteration #2”. Within the first phase (iteration #1), the intake air volume requirements were determined by Genivar Engineering based upon the mine’s 18-month production schedule, from January 2006 to June 2007. This was then extrapolated through to year 2018 based upon an “airflow to production tonnage” ratio. Within the first phase, the inactive levels of the mine were considered as being closed (unventilated), with no airflow allocated to these levels. However, an additional global allowance of 20% was allocated to account for leakage into these inactive areas (see section 5.4.2 – Iteration #1). Within the second phase (iteration #2), CANMET-MMSL has identified that a major deficiency in iteration #1 was an increased potential for uncontrolled recirculation within the mine’s auxiliary ventilation system. Further to this CANMET-MMSL has determined that in order to avoid uncontrolled recirculation within the mine’s auxiliary ventilation system, a minimum airflow velocity of 0.25 m/s needs to be maintained past the auxiliary fans. Consequently, in order to avoid uncontrolled recirculation and at the same time to  40  accommodate other operational needs, the mine’s initial intake air volume determined through iteration #1 was re-evaluated by CANMET-MMSL (see section 5.4.2 – Iteration #2). Airflow Calculation Based Upon the Mine’s Electronic Database (3rd Stage Review) At the Kidd Creek Mine, at the end of every shift, all mining activities are recorded into an electronic database referred to as SIMS by the shift supervisors and equipment operators. These data detail where production diesels were operating during the production shifts, their location and for how long. These data was imported into MSExcelTM, it was then processed into air volume requirements based upon the engine size and the number of vehicles that were operating on the production levels. Despite being complex, labour intensive and time consuming these two conventional techniques did not agree. Details with respect to these two case studies are presented in sections 5.3 and 5.4  3.3  Mining Process Model Development Using AutoModTM  This thesis introduces a novel ventilation design concept which enables the determination of a mine’s “true” intake air volume requirement through discrete-event mining process simulations performed on a mine model that integrates production operations, secondary mining operations and fan activities. Through this innovative method the accuracy of a mine’s activity based airflow requirement would only depend upon the detail to which an underground operation is modeled and the accuracy of the mining activity data entered by the user (i.e. equipment size, fan duties, activity durations). To determine the activity based air volume requirements, a generic “Mining Process” model was developed in collaboration with Penguin Automated Systems. Based upon the mining activity schedules and fan operating duties entered by the user, the ventilation output data generated by discrete-event process simulations was automatically processed into “traditional” as well as “activity based” air volume requirements. This information was then used to determine the “life-cycle” airflow demand of a mine (Mine A). For practical purposes and to accurately determine the economic and environmental benefits of this new ventilation design concept, the life-cycle airflow demand schedules for traditional and activity based ventilation were determined for an existing mine, namely “Mine A”, a  41  large and deep metal mine, employing the vertical retreat mining method to extract the ore reserves.  3.4  Ventilation System Design Based upon the Mine’s Life-Cycle Airflow Demand Schedule  To determine Mine A’s energy consumption for both traditional and activity based ventilation practices, firstly, its primary and auxiliary ventilation systems needed to be solved and balanced. This was performed through ventilation simulations performed on Mine A’s primary and auxiliary ventilation systems using VnetPCTM. The primary ventilation system of “Mine A” was developed according to mine layouts and information provided by the mine’s engineering department. The technical characteristics of the mine surface and booster fans were also provided by the mine’s engineering department. Mine A’s life-cycle airflow demand schedule showed that within certain time periods of its operating life, the intake air volume requirements varied according to the planned production targets. Consequently, for both traditional and activity based airflow requirements, Mine A’s primary ventilation system was solved & balanced and its fan power and ventilation operating costs determined for the following time periods: years 2008-2010, 2011-2013, 2014-2017 and 2018-2020.  3.5  Economic and Environmental Benefits  Based upon the output data generated though ventilation simulations, Mine A’s energy consumption was determined for traditional and activity based ventilation practices. The savings in energy consumption generated by an activity based ventilation system were transformed in savings in greenhouse gas (GHG) emissions expressed in tonnes of equivalent CO2.  42  4.  UNDERGROUND OPERATIONS SURVEY  4.1  Mining Methods Employed by Canadian Mines  The mining methods employed by the Canadian underground operations can be grouped in the following categories: •  Unsupported methods: - Room-And-Pillar Mining - Stope-And-Pillar Mining - Shrinkage Stoping - Sublevel Stoping (ring drilling/parallel drilling) - Vertical Retreat Mining (VRM)  •  Supported methods: - Cut-And-Fill Stoping  •  Caving methods: - Longwall Mining - Sublevel Caving - Block Caving  •  Narrow vein methods - Cut-And-Fill (Narrow vein) - Longhole (Narrow vein)  The unsupported class consists of the methods in which the rock is basically self-supporting and for which no major artificial support is necessary to carry the load of the overlying rock. Of the unsupported mining methods, room-and-pillar and stope-and-pillar mining employ horizontal openings and low opening-to-pillar ratios. Shrinkage and sublevel stoping methods usually employ vertical or steeply inclined openings and high opening-to-pillar ratios [Hartman & Mutmansky, 2002]. The supported methods are those methods that require some type of backfill to provide a substantial amount of artificial support to maintain stability in the extraction openings of the mine. The supported class is usually employed when the other two categories of methods, unsupported and caving are not applicable. The caving class includes the mining methods in which the extraction openings are designed to collapse; that is caving of the ore or rock is intentional and is the very essence of the 43  method. The sublevel and block caving methods have application to inclined or vertical, massive deposits, almost exclusively metallic or non-metallic (Hartman & Mutmansky, 2002).  4.2  Identification of Canadian Mines that Would Benefit the Most from an Activity Based Ventilation Control System  To evaluate economical and environmental benefits of this new ventilation design concept, all mining methods employed by Canadian mines were reviewed and assessed. A brief description of the extraction methods, their characterization and applicability to various geological, geotechnical and mining conditions was presented in the previous section. Mine ventilation studies [Hardcastle et al, 2006] [Kocsis et al, 2003] showed that an important factor during the design process of the primary and auxiliary ventilation infrastructure is the employed extraction method(s). This is because the extraction method can ultimately dictate the layout of the development and production workings, ground control techniques, ore/waste transportation systems, thus the size and number of production and service equipment. An initial classification of the Canadian mines can be based upon the most frequent extraction methods employed. For example, Table 3 shows that the most utilized mining method in Canada is sublevel stoping, which is presently employed by 28 operations of which 12 base metal and 16 precious metal mines, followed by cut-and-fill (13 operations), vertical retreat mining & shrinkage stoping, room-and-pillar, continuous mining, narrow vein mining and sublevel caving. This table was compiled based upon a survey performed and published by the Canadian Mining Journal’s 2005, 2006, 2007 and 2008 Mining Sourcebook. It should be mentioned that not every operation responded to the questionnaire sent by the Canadian Mining Journal and only those underground operations who responded were taken into account in this survey. Based upon the geological conditions, geotechnical properties of the mineral deposit and operating characteristics of the mining process, the Canadian underground operations can be grouped in two main categories, as follows: •  “Hardrock” mines, which include the base metal, precious metal, uranium and diamond mines with a discontinuous cyclical mining process, and  •  “Softrock” mines, which include the potash, gypsum, salt and coal mines, with ore reserves extracted through a continuous mining process 44  In the following two sections for the mines that were grouped in the above mentioned categories (i.e. hardrock/softrock), the applicability and the potential economic and environmental benefits of this new ventilation design concept will be evaluated and discussed. Table 3: Classification of the Canadian mines based upon mining method utilization frequency Mines Hardrock Mines Softrock No. Mining Method Employing Mines this Method Base Precious Uranium Diamond Potash Salt Metal Metal 1 Room-and-Pillar 5 (7%) 3 1 0 0 0 1 2 Stope-and-Pillar 2 (3%) 1 1 0 0 0 0 3 Shrinkage Stoping 8 (11%) 2 6 0 0 0 0 4 Sublevel Stoping 28 (37%) 12 16 0 0 0 0 5 Vertical Retreat Mining 8 (11%) 7 0 0 1 0 0 6 Cut-and-Fill Stoping 13 (17%) 5 6 1 0 1 0 7 Sublevel Caving 3 (4%) 2 0 0 1 0 0 8 Block Caving 0 (0%) 0 0 0 0 0 0 9 Narrow Vein Mining 3 (4%) 1 2 0 0 0 0 10 Continuous Mining 5 (7%) 0 0 0 0 5 0  4.2.1 Hardrock Mines – Base and Precious Metal Mines Hardrock mines can be further divided in two categories; the first category includes the “highly mechanized” mines that extensively use diesel-powered equipment during the development and production operations as well as for personnel transportation, and “nonmechanized” mines, that still employ hand-operated compressed air equipment. Due to the dynamic nature of the development and production workings constantly redefining where ventilation is needed and the cyclical nature of the mining process (drilling-loading-blastingmucking), hardrock mines usually maintain more production stopes available (active) as it would be required by production rates. Usually, in “hardrock” mines, the primary ventilation design criteria are the dilution and the removal of potentially harmful by-products generated by the mining activities. This can include respirable dust generated during the ore/rock breaking and handling activities; respirable particles created in the combustion chambers of the diesel equipment (i.e. carbon soot); gasses liberated from the strata or generated by the diesel-powered mining equipment; radio-active gasses or attachments to particles; fumes from detonation of explosives in the 45  production workings; and heat resulting from the strata, auto-compression and mining equipment. The removal of blasting fumes, although important, is only a short-lived periodic concern; here the primary consideration is that it is performed in a timely manner in order to avoid production delays. Consequently, the hard rock mines where radiation, diesel exhaust dilution and heat removal are the most important concerns have the highest ventilation requirements. Furthermore, due to the fact that the hardrock mines maintain a large number of production stopes active, they can be significantly over-ventilated. As a result, the greatest opportunity to reduce underground fan power and consequently their energy consumption is within the Canadian hardrock mines. An activity based ventilation system may not be beneficial for uranium mines, where a constant need for ventilation is required to dilute and remove the radioactive elements that are continually produced regardless of the presence or absence of the mining equipment. With regard to Canada’s diamond mines, for the most part these are still surface operations, and where they have started to move underground their ventilation energy usage is minor compared to the deep and heavily mechanized metal mines. It can therefore be concluded that this new ventilation design concept based upon the “life-cycle” airflow demand schedule would be most beneficial for the hardrock mines especially for the deep and highly mechanized metal mines  4.2.2 Softrock Mines In the Canadian “softrock” mines, which include the potash, gypsum, salt and coal mines, the ore reserves are usually extracted through a continuous mining process. In these mines the ore at the face is usually fragmented and transported to the surface using electrically-powered equipment. As a result, the need to immediately dilute blasting fumes and diesel exhaust fumes is not an issue. The Canadian softrock mines are in general operating at shallow or medium depth and the electrical mining equipment is approximately 3 times more efficient than diesel equipment. Consequently, the need to control the heat generated during the mining processes does not present the same importance as in deep metal mines. Furthermore, in softrock mines (with the exception of coal mines), dust is generally classified as a temporary issue and compared to hardrock mines has a low order of concern.  46  In softrock mines, diesel powered equipment is only used for personnel transportation and the air volumes required by these smaller diesel engines are significantly less than the air volumes required by large production equipment within the metal mines. Where the ore reserves are extracted using diesel equipment (i.e. salt mines), the intake air volume can be significant, however, during the non-productive time periods these air volumes cannot be reduced due to the large cross-sectional areas of the production and haulage openings. In other words, if the intake air volume in a salt mine would be reduced according to activity based ventilation needs, the airflow velocity along such large openings would fall below the minimum value required by health and safety regulations [Hardcastle et al, 2006]. So, it can be concluded that in softrock mines, this new ventilation design concept would not provide the same economic and environmental benefits as in hardrock mines.  47  5.  Case Studies: COMPARISON OF CURRENT VENTILATION DESIGN CRITERIA WITH THE NEW DESIGN CONCEPT IN DEEP METAL MINES  5.1  Introduction  During the operating life of an underground operation there will be periods when substantial changes to the ventilation system need to be undertaken due its dynamic nature. This usually occurs when: •  The mine is planned to deepen or a new mining zone is to be opened,  •  Production workings become sufficiently remote from surface connections such that the mine’s existing intake infrastructure can no longer provide the required air volumes to the production areas economically, and  •  Two or more mining zones with individual intake and exhaust systems are to be interconnected  In Canada, the need for some mines to re-evaluate their ventilation system and extraction methodology is becoming extremely important in the light of increasing mining depth, escalating energy prices, the environmental implications associated with energy consumption and continuing pressure to remain competitive. This challenge to the industry was shown in a study funded by the Deep Mining Research Consortium (DMRC), which indicated that in Canada, with increasing mining depth and unchanged operating practices, underground ventilation and ore transportation would experience a significant escalation in both capital and operating costs [Campbell, 2005]. With respect to minimizing ventilation costs, the general options are either changing the ventilation design drivers, such as the need to dilute diesel exhaust, or improving the utilization efficiency of the air supplied to the production workings. To move away from the diesel exhaust design criteria, such options as tele-remote mining from surface and the introduction of fuel cell powered production equipment have both been considered. Furthermore, in order to improve the efficiency of a mine’s ventilation system other technologies such as controlled recirculation and demand-based delivery systems have also been tried. However, in order for an underground operation to embrace and implement a new technology that would significantly change the way the mine was previously ventilated, this  48  new technology would need to be evaluated technically and proven economically beneficial [Hardcastle et al, 2006].  5.2  Ventilation Redundancy and its Importance in Measuring the Efficiency of Underground Ventilation Systems  An important phase of the ventilation system optimization exercise is to determine the ventilation redundancy within the primary and auxiliary systems. Mine ventilation design has improved significantly with the use of network simulation programs. However, it should be mentioned that the current network simulation programs do not perform any creative design work and their use in the design process is normally based upon achieving a desired airflow at specific times in a mine’s life. One limitation of this design approach is that the results are usually based upon peak ventilation demands that assume production activities as being continuous. As a result, there can be a significant and costly ventilation redundancy in the solution. This is because most production workings employ discontinuous mining processes and they do not need to be continuously ventilated. However, the determination of ventilation redundancy in large and deep metal mines can be complex, problematic and extremely time consuming. To highlight these, in sections 5.3 and 5.4 various airflow evaluation methods that were applied to determine ventilation redundancy in deep metal mines as well as a summary of findings from two case studies will be presented, as follows: •  Research study that explores the potential benefits of introducing ventilation automation at Vale Inco’s Creighton Mine  •  Feasibility study analyzing ventilation utilization efficiency at Xstrata’s Kidd Creek Mine  5.3  Case Study #1: Determination of Ventilation Redundancy at Creighton Mine – Vale Inco  This feasibility study [Hardcastle, S.G. & Kocsis, C., 2004] explored the potential savings to be gained with the introduction of ventilation controls within the mine’s primary and auxiliary ventilation system. The focus of this study was to determine for what duration it was necessary to provide ventilation to the main production functions, namely drilling, blasting clearance, mucking, dumping and backfilling. The findings were then used in 49  association with ventilation simulation techniques to indicate what order of ventilation savings may be achieved. Creighton mine in the Sudbury basin, Vale Inco’s oldest operation, was discovered in 1856 and made its first ore shipment in 1901. Over its +100 years of production, many mining methods have been used, including open pit, shrinkage, blasthole, panel caving, square set, undercut-and-fill and vertical retreat (with blastholes). Changes in respect to the employed mining methods came as a function of various operating factors such as the introduction of mechanization and new ground control and heat management methods as the mine deepened. Today, production relies upon mechanized vertical retreat mining (VRM) with delayed backfill and is concentrated below 1,500m level [Hardcastle & Kocsis, 2004]. At Creighton mine, installing widespread monitoring, control and communication systems for ventilation purposes alone would be very expensive and could only be justified if such a highly controlled system would generate measurable and proven cost benefits. Consequently, the first objective of this study [Hardcastle & Kocsis, 2004], was to predict how the operating costs would increase as the production and development workings deepen if the mine uses the existing ventilation management practices. It then explored the potential cost savings with the introduction of a ventilation-on-demand based system through temperature and gas monitoring activity studies.  5.3.1 Air Volume Requirements at Creighton Mine For Creighton mine, as with other deep mines, heat mitigation is increasingly becoming more dominant than diesel exhaust dilution in the overall design requirements of the primary ventilation system. Within the production areas, the air volumes by the local auxiliary ventilation system are still based upon Vale Inco’s standard design criteria of 0.079 m3/s/kW (125 cfm/bhp) for diesel emission dilution [Stachulak, 1989], which is higher than the local statutory requirements of 0.063 m/s/kWh (100 cfm/bhp) to account for leakage. With respect to the specific flows needed at each production level, Table 4 lists the estimated virgin strata temperature, the expected intake wet-bulb air temperature and the airflow requirements as a function of the daily production rate and mining depth. The strata temperatures are based upon a datum virgin rock temperature of 43.3 0C at 2,195 m depth and a geothermal step of 1 0C per 55 m. The intake air temperatures were derived through climatic modelling and the calculated air volumes take into account the heat generated from 50  the mining machinery, the geothermal gradient and the mine’s production rate. Of note, here, is the significant difference between the virgin strata temperature and the intake air wet-bulb temperature. This is the result of the mine using a natural heat exchanger comprised of broken rock in the intake air system [Stachulak, 1989]. This mass of broken rock through its annual cooling and heating cycle, respectively heats sub-zero intake air in winter and then subsequently cools the warm air entering the mine during summer. The net result at 250m below surface is an air temperature of 2.8 0C +/- 1.7 0C year round. However, during summer this effectively free source of cooling has a limited capacity. Table 4: Temperature and air volume requirements as a function of mining depth at Creighton Mine - Vale Inco [Stachulak, 1989] Depth Below Surface Datum (H) m ft 2,134 7,000 2,195 7,200 2,256 7,400 2,335 7,660 2,371 7,780 2,408 7,900 2,444 8,020 2,481 8,140 2,499 8,200  Virgin Strata Intake Air Wet-Bulb Temperature (VRT) Temperature (TWB) (0 C) (0 F) (0C) (0 F) 16.35 61.43 43.3 110.0 16.58 61.85 44.4 112.0 16.67 62.00 45.9 114.6 17.11 62.80 46.6 115.8 17.41 63.34 47.2 117.0 17.72 63.90 47.9 118.2 18.01 64.42 48.6 119.4 18.33 65.00 48.9 120.0 18.56 65.40  Required Air Volume (Q) m3/s/tpd cfm/tpd 0.099 209 0.106 225 0.114 241 0.128 272 0.135 286 0.142 300 0.148 314 0.155 328 0.170 361  Table 4 also shows that the required air volumes at the production workings will significantly increase as the mine deepens. For example, between 2,134m and 2,481m, for an increase in vertical distance of 347 m or 16%, the predicted air volume needed for heat mitigation increases disproportionably by 57%. Such increases have two effects; they will cause the ventilation operating costs to dramatically increase through the use of more electrical power and also accelerate the depletion of the natural heat exchanger’s cooling capacity. As the natural source of cooling has a finite capacity, the tonnage that could be mined during summer may become increasingly limited as the production workings deepen. Alternatively, additional mechanical cooling, at considerable capital and operating costs may need to be considered. However, another possibility is that the mine explores methods to conserve the natural heat exchanger’s cooling potential through an efficient air management system. To deliver the required air volumes to the development and production workings, a total intake air volume of 645 m3/s is needed. From this approximately 170 m3/s is being sent 51  down the No. 9 shaft, a dedicated main intake airway through which fresh air is delivered to the main levels of the upper orebody (3800’ to 7000’ levels). Furthermore, approximately 280 m3/s and an additional 195 m3/s is delivered down the main fresh air raise (MFAR) and the 3rd fresh air raise system (FARS), respectively, to the deep orebody’s development and production workings [Kocsis & Hardcastle, 2004].  5.3.2 Mine Activity Monitoring Through Temperature Studies Creighton mine utilizes 13 primary fans within its primary ventilation system, namely five sets of two fans in parallel acting as boosters distributed through its underground intake system, and three fans in parallel as surface exhaust. The total combined rated power of these fans is 11,222 kW. Table 5, derived through ventilation simulations shows the net effects of increased air volumes with depth and changes in flow and power consumption due to the concentration of production areas at depth through four planned stages of development (between 2002 – 2019). Table 5 shows that although the mine’s total intake air volume increases by 7%, from 626 m3/s to 670 m3/s, the flow supplied to the deep orebody increases more dramatically, firstly from 150 m3/s to 315 m3/s as it comes fully on-line, and then to 447 m3/s with additional depth. Consequently, the associated power requirements of the primary system will increase by 45% [Hardcastle & Kocsis, 2004]. Table 5: Increasing airflow and power needs at Creighton Deep [Hardcastle & Kocsis 2004] Time Frame  Deep Orebody Activity  Conceptual Schedule  Base Model  1 Production Level, 1 Development Area 3 Production Levels, 2003-2007 1 Development Area 3 Production Levels, 2007-2011 1 Development Area 3 Production Levels, 2011-2015 1 Development Area  Levels  Tonnage Airflow Power Cost Mine Deep Primary Auxiliary Primary Auxiliary Total Mine Deep Orebody Orebody (tpd) (tpd) (m3/s) (m3/s) (kW) (kW) ($MCan) ($MCan) ($MCan)  7400 to 4000 7530 7400 to 4000 7720 7720 to 7900 7900 to 8080 8080 to 2015-2019 3 Production Levels 8200  1300  626  150  5,559  3,200  2.42  1.40  3.82  1700  634  315  5,775  3,360  2.52  1.46  3.98  1500  645  382  6,112  3,528  2.66  1.54  4.20  1500  662  433  6,811  3,704  2.97  1.61  4.58  1500  670  447  8,082  3,890  3.52  1.69  5.21  At the time of the analysis, Creighton mine employed approximately 200 auxiliary fans, with motor sizes ranging from 37 to 56 kW. However, it was difficult to calculate their operating costs due to lack of data as to when any of these fans were operating. For example, using an average of 46.5 kW per fan, the total installed power of the auxiliary system would be in the 52  order of 9,300 kW. However, mine data showed a total connected load for the auxiliary system of only 6,200 KW. This indicated that on average only 130 of these auxiliary fans were operating at any time. Furthermore, if this connected load was continuous, the annual operating cost would be $2.7M, but the mine reports a cost of $1.4M [Vale Inco, 2004], which tends to indicate that on average only 70 of these auxiliary fans were operating at any time with an average connected load of 3,200 kW. Overall, this study showed that in order to accurately determine the power consumption within the mine’s auxiliary ventilation system a more detailed evaluation was needed. Consequently, to calculate the auxiliary requirements shown in Table 5, an increase of 5% per stage was assumed starting from 3,200 kW, which represented the base model condition supplied by the mine [Hardcastle & Kocsis, 2004]. Temperature Studies At the time of the mining activity study, Creighton mine’s operating schedule was two 10hour back-to-back shifts per day followed by a 4-hour clearance break, and of the 14 shifts per week, two week-end shifts were designated as maintenance and service shifts. The activity within the production shifts and general need for ventilation to be supplied to specific mining areas was determined by CANMET-MMSL [Hardcastle & Kocsis, 2004] through a combination of temperature and gas monitoring. During 2001 and 2002, the following four mining activities were monitored with temperature and humidity sensors for a total of 125 days: •  Two LHD drawpoints for 28 and 33 days,  •  One LHD backfilling site for 28 days,  •  One LHD ore-pass site for 8 days, and  •  One long-hole drilling site for 26 days  Where possible, the temperature and humidity sensors were installed inside the auxiliary ducts to monitor intake air conditions, within the draw-point working area and along the return airways to monitor the conditions of the return air. The locations of the temperature and humidity monitors are shown in Figure 6, as follows:  53  Figure 6: Location of temperature and humidity monitoring sensors on 7400L – mucking operation [Hardcastle & Kocsis, 2004]  No. 2 Data Logger Fresh Air - Auxiliary Duct Discharge  No. 1 Data Logger Fresh Air - Auxiliary Duct Intake  No. 3 Data Logger 4360 Stope Area  No. 4 Data Logger 2nd Access Return  No. 5 Data Logger 4100 X-Cut Return  This monitoring program was able to identify when activities were taking place, plus the climatic conditions generated by the mining equipment within the production workings. For example, Figure 7, a sample of this monitoring program shows the temperature conditions at five locations around a mucking operation. In general the temperature trends at all locations agree, showing to varying degrees when activity started and ceased. When such data is studied on an expanded time scale it is even possible to discern the cycle-time of the production equipment, and also how long it typically takes for temperature conditions to return to the drawpoint background conditions. Figure 8, which represents the same mucking operation but on an expanded scale, shows an average load-haul-dump cycle of 7 to 8 minutes, and that 10 to 15 minutes are required for the majority of the equipment’s temperature influence to dissipate. From this type of monitoring and based upon the temperature conditions in the working area, it was found that the auxiliary ventilation system was only needed to operate continuously from the arrival of the mining equipment, through its active period and up to 10 minutes after its last departure [Hardcastle & Kocsis, 2004]. 54  Figure 7: Typical dry-bulb temperatures plot around a 250 kW LHD moving ore from a draw-point to the orepass [Hardcastle & Kocsis, 2004] Creighton 7400 - Heat Study  Wednesday 28/11/01 37 BLAST CLEARANCE 36 35 34  Intake has lowest temperature during activity, but influenced by activity. Duct discharge 2C higher - higher than exhaust during active periods. Higher than all during non active. Both Accesses active -show independant effects. Activities do not raise general air temperature by above duct discharge, exhaust is also cooled by excess air delivered to level during non-active periods. Activities raise face temperature by up to 3C above discharge Area Exhaust can be lower than intake, due to loss of fan heat or local evapouration. evaporation  33 Temp. (C)  DAY SHIFT  NIGHT SHIFT  Mucking in 2nd Access  Mucking in 1st Access  32 31 30 29  288 mins mucking - 33 of 35 buckets 12:01  28  08:55  01:08  10:32  467 mins break 27 22:00  0:00  2:00  4:00  6:00  13:30  13:12  89 mins 18 mins 8:00  10:00  12:00  341 mins mucking - 37 of 43 buckets  14:00  15:30  18:08  21:37 22:08 31 mins  158 mins break 16:00  18:00  20:00  22:00  00:20 450 mins break 0:00  Time (hh:mm) Intake Raise  Duct Discharge  Face 1st Access  Return 2nd Access  Area Exhaust  2:00  4:00  55  Figure 8: Expanded dry-bulb temperature plot around a 250kW LHD showing cycle times [Hardcastle & Kocsis, 2004] Creighton 7400 - Heat Study  09:00  12:05  10:19  15:10  13:29  12:55  40 39 38  33  37 36  31  35 29  11 Mucking Cycles - average 7.90mins  14 Mucking Cycles - average 7.77mins 8 Mucking Cycles - average 7.14mins  34 33  27  32 31 Overall 33 Mucking Cycles - average 7.67mins  25 8:45  9:15  9:45  10:15  10:45  11:15  11:45  12:15  12:45  13:15  13:45  Time (hh:mm) Return 2nd Access  Area Exhaust  Face 1st Access  14:15  14:45  15:15  30 15:45  1st Access Face Temp. (C)  2nd Access & Area Exhaust Temp. (C)  Wednesday 28/11/01 35  56  Based upon such temperature monitoring it was possible to obtain a long-term “mucking activity log”, as shown in Figure 9. Here it can be seen that the activities at a mucking location are far from continuous during the 767 hours of monitoring time period. Sixty periods of LHD activity,  totalling  123.7  hours  were  identified,  along  with  another  32.6  hours  miscellaneous/mucking activity in the immediate area. On allowing for temperature clearance, the auxiliary ventilation system was only required to operate 167.4 hours or 21.8% of the monitored time; this is also a measure of the ventilation system’s utilization. Figure 9 also shows the variability of mucking activity, namely, between Jan. 06 – Feb. 09, 2002, there were 19 shifts with no activity. Furthermore, on the day shifts activity could start as early as 8:24AM or as late as 2:30PM and then end as early as 10:30AM or as late as 3:50PM. Also, the active periods in any day could be as short as 45 minutes and as long as 358 minutes. Breaks in activity ranged from 40 minutes through to 161 hours, plus there was no apparent consistently scheduled mid-shift meal breaks [Hardcastle & Kocsis, 2004]. With respect to possible ventilation control, for this location, the primary ventilation system would only have to operate at its maximum level for two periods of 7.5 hours per day, 6 days per week to cover all potential operating scenarios (i.e. 56% of the regular working week). During the remaining period, it could operate at a reduced level. Based upon the temperature trends, similar activity logs (as shown in Figure 9 for mucking), were carried out for drilling, backfilling and ore dumping. These activity logs indicated firstly, the periods of time where the primary ventilation system could be reduced and secondly, the overall operational requirements for the auxiliary system within such production areas.  5.3.3 Production Blast Monitoring Longhole blasting was also captured during the mining activity monitoring with the temperature sensors, however this only provided information on their scheduling. To determine blast gas clearance times, three blasts were monitored at four locations using Draeger Multi-WarnTM gas monitors. A typical sample of the blast monitoring is shown in Figure 10.  57  Figure 9: Activity log derived from temperature monitoring at a mucking draw-point operation [Hardcastle & Kocsis, 2004] Mucking Activity Blast Clearance  Day Shift  Night Shift  BLASTS  6-Jan Monday  8-Jan 10-Jan  Start of Monitoring  Tuesday Wednesday  12-Jan  Thursday Friday Saturday  14-Jan  Sunday Monday  15:50  1:07  41.8 hrs  18-Jan  Start 8:24AM  Wednesday Thursday Friday Saturday  20-Jan  Sunday Monday  22-Jan  Tuesday Wednesday  24-Jan 26-Jan  Thursday Friday Saturday  28-Jan  Sunday Monday  30-Jan  End 3:50PM  4.9 hrs  Start 2:30PM Miscellaneous Activity  47.1 hrs  50.1 hrs  End 10:30AM Stope was monitored for 767 hrs - Activity apparent from two sources During 31 days (62 shifts) - 60 periods of "mucking" & 21 other identified - 19 shifts no activity  55.4 & 78.8 hrs  Active mucking time 123.7hrs (16.1%) - Other activity 32.6 (4.3%) Tuesday Wednesday Active mucking time plus 60 x 10mins for temp. clearance becomes 133.7hrs (17.4%) Thursday Inactive periods between "shifts", 1:20-8:20 & 16:00-17:40, 277.3hrs (36.2%)  1-Feb  Friday Saturday  3-Feb  Sunday Monday  5-Feb  Tuesday Wednesday Thursday  7-Feb  2.4 hrs  8:24  Tuesday  16-Jan  1.9 hrs 17:43  16.2 hrs  Active period during "shifts" plus 60 x 10mins for temp. clearance vs potentially active (34.2%)  66.0 hrs  End of Monitoring  Friday  9-Feb 3:30  161.2 hrs  5:30  7:30  9:30  11:30  13:30  15:30  17:30  19:30  21:30  23:30  1:30  3:30  58  Figure 10: Sample of blast contaminant decay monitoring [Hardcastle & Kocsis, 2004]  400  8000 BLAST 02:15 CO2 TLV = 5000ppm Time to 1/2 TLV= 02:15 to 02:30  300  NO TLV = 25ppm Time to 1/2 TLV= 02:15 to 02:37  250  NO2 TLV = 3ppm Time to 1/2 TLV= 02:15 to 02:44  200 150  CO TLV = 25ppm Time to 1/2 TLV= 02:15 to 03:02  100 50 0 2:00  7000 6000 5000 4000 3000 2000 1000  2:30  3:00  3:30  4:00  Time CO  NO  NO2  CO2  4:30  0 5:00  CO2 Concentration (ppm)  Gas Concentration (ppm)  350  59  With the exception of carbon dioxide, during the VRM blast, all other sensors went off-scale. However, on excluding this region and upon re-scaling, it was found that all the monitored gasses including carbon dioxide, carbon monoxide, nitrogen dioxide and nitric oxide displayed the same decaying profile. Throughout the blast clearance evaluations, carbon monoxide was always the last gas to clear to 50% of its time weighted average limit (TWAEV). Consequently, it would be the most reliable gas to be monitored in order to provide safe working conditions in the production area. Due to the variable conditions of the VRM blasts, off and unknown auxiliary fan conditions, the VRM blast clearance results were inconsistent. However, the study showed that the gas clearance could be as short as 35-40 minutes as opposed to the mine’s current 4–hour clearance window (e.g. between shifts), plus the allowance in the shifts.  5.3.4 Determination of Ventilation Redundancy at Creighton Mine This case study showed that if the Creighton mine would continue with its current strategy of providing ventilation continuously, the power requirements and costs would increase significantly. In addition there is the potential for the mine’s natural cooling capacity to become insufficient. This presented a concern, not only due to the escalating ventilation costs that could affect mining sustainability, but also for the potential environmental impact of increased green house gas (GHG) generation. Through temperature, blast clearance and activity monitoring, it has been shown that the flow through the primary system could be reduced to a lower operational level 44% of the time. This monitoring has also shown that on average an auxiliary fan in the production area may be required as little as 20% of the time. With the appropriate control of the primary and auxiliary fans, it has been shown that through “time-of-day” and “activity” based controls, the mine’s annual power requirements and hence operating cost can be reduced by 30-40% compared to an extrapolation of their current “constant” delivery system. Table 6 shows that for the 2003-2007 and 2015-2019 periods, the net effect of the management of flow, namely flow reductions during non-productive periods can reduce the mine’s total intake air volume by 18% and 11%, respectively. Furthermore, these will result in average flows through to the year 2019 that are less than those currently supplied. Hence the need for introducing mechanical refrigeration can be avoided.  60  Table 6: Combined ventilation operating costs/savings with time-of-day and activity based controls [Hardcastle et al, 2006] Time Depth Period  Operating Condition  (m) 24/7 2003 to 2007  2250 to 2350  2015 to 2019  2450 to 2500  Average Net Flow Primary Reduction (%) Flow (m3/s) 634  Time-of-Day Time-ofDay/Activity 24/7  517 670 599  Average Power Auxiliary Combined System (kW) (kW) 3,360 9,135  Total Cost $M(Can)  Net Savings $M(Can) (%)  3.98  3,869  2184  6,053  2.64  3,869  1210  5,079  2.21  8,082  3,890  11,972  5.21  5900  2528.5  8,428  3.67  5900  1400  7,300  3.18  18%  Time-of-Day Time-ofDay/Activity  Primary System (kW) 5,775  11%  1.34 34% 1.77 44% 1.54 30% 2.03 39%  This case study was successful in showing ventilation redundancy within Creighton mine’s ventilation system that was based upon constant flow delivery. However, determining the ventilation redundancy within the auxiliary ventilation system was difficult, which required assumptions to be made. Another significant disadvantage of this conventional analysis method is that it is extremely labour intensive and time consuming. To accurately determine the required air volumes in the specific mining areas of this complex multilevel operation, the mining activities needed to be monitored for a prolonged period of time, namely for a total of 125 days. Furthermore, to identify when activities were taking place plus the climatic conditions within the production workings, the temperature and gas monitoring units were programmed to collect climatic data at 1-minute and 30-second intervals, respectively. These monitoring units generated a very large amount of electronic data, which then needed to be processed and compiled in temperature and blast contaminant trends that were able to show when activities started or ceased as well as the climatic conditions they created.  5.3.5 Prediction of Cost Savings through the Elimination of Ventilation Redundancy at Creighton Mine The activity logs have shown that maximum primary ventilation is required 56% of the time in a standard working week. For the remaining 44% of the time, it could be reduced to a lower rate. The relative benefit of this reduced operational period is very dependant upon the minimum flow specified. In the cost reduction analysis, it has been assumed that all flows in productive areas fall back to a volume suitable for service vehicles. Table 7 compares the results of ventilation simulations based upon full and reduced flow requirements for the 61  original schedule provided for the 2003-2007 and 2015-2019 periods. This table shows that as the mine deepens, the net reduction in flow decreases, as more areas have to be ventilated. Despite this, the cost of running the primary ventilation system could be reduced by on average 30%, and the majority of the operating cost (>75%) would be attributable to the full ventilation period. Table 7: Primary ventilation system operating cost/savings with time-of-day controls [Hardcastle et al, 2006] Time Period  Working Depth (m) 2250 to 2350 2450 to 2500  2003 to 2007 2015 to 2019  Operating Condition Maximum Minimum Reduction Maximum Minimum Reduction  Mine Airflow (m3/s) 634 369 42% 670 509 24%  Deep Airflow (m3/s) 315 137 57% 447 280 37%  Primary Power (kW) 5,775 1,575 73% 8,082 3,297 59%  Annual Operating Adjusted Cost Time Cost ($MCan) Factor ($MCan)  Total Cost ($MCan)  54% 46%  1.68  2.52 0.69 73% 3.52 1.44 59%  54% 46%  1.36 0.32 Net Saving 1.90 0.66 Net Saving  33% 2.56 27%  Determining the mine-wide cost savings for the auxiliary system is more problematic. Unlike the primary system, the auxiliary fans would be either on, or off. The activity logs have shown that auxiliary fans may be, on average, only required 20% of the time. Furthermore, according to the mine, most of the operational fans are in the production areas and could be subject to some form of time/activity based control. In Table 8, it has been assumed that 80% of the installed auxiliary fan power would be controlled, that it would firstly operate 56% of the time in-line with the primary system, and then further optimized to operate only 20% of the time. The net result in operating cost is a 35% reduction with shift-based time-of-day controls, and a 64% reduction with activity-based controls. Table 8: Auxiliary ventilation operating costs/savings with time-of-day and activity based ventilation controls [Hardcastle et al, 2006] Time Depth Period (m) 2003 to 2007  2250 to 2350  2015 to 2019  2450 to 2500  Operating Condition  Auxiliary Fans Annual Operating Adjusted Total Net Total Control Controlled Cost Time Cost Cost Saving Power Factor Power ($MCan) Factor ($MCan) ($MCan) (kW) (kW) Fixed 20% 672 0.29 100% 0.29 35% Time-of-day 3,360 56% 0.66 0.95 80% 2688 Controlled 1.17 Activity 0.53 64% 20% 0.23 24/7 Total 1.46 Fixed 20% 778 0.34 100% 0.34 35% 56% 0.76 1.10 Time-of-day 3,890 80% 1.36 Controlled 3112 Activity 0.61 64% 20% 0.27 24/7 Total 1.69  62  Table 6 combines the cost benefit analyses of the primary and auxiliary fan systems. However, as both independent analyses are based upon certain assumptions, the final savings only serve to give an indication of the potential savings, as follows: •  Simple “time-of-day” controls on both the primary and auxiliary fans could produce at least a 30% reduction in power consumption and operation cost  •  Simple “time-of-day” controls on the primary fans and “activity-based” control of auxiliary fans could produce at least a 39% reduction in power consumption and operating cost.  This table also shows that the average primary airflow with time-of-day based controls is less than the 626 m3/s supplied in the initial model. Consequently, if the mine’s natural heat exchanger retains the same capacity and is currently satisfactory; it should be able to maintain the same conditions throughout the development of the Deep Orebody plus be able to support other areas.  5.4  Case Study #2: Ventilation Utilization Efficiency at Kidd Creek Mine – Xstrata PLC  In 2005, the Kidd Creek mine initiated three stage review of its ventilation system to minimize potential ventilation redundancy and increase the efficiency of the ventilation system. This was considered necessary, in an attempt to reduce the mine’s current and future power consumption and its associated ventilation costs. The first stage review consisted of a brainstorming forum organized by the mine’s engineering department. In the second stage, the mine’s intake air volume requirement was determined by Genivar Engineering based upon future production planning. Further to this, Kidd Creek mine requested CANMETMMSL to review the work performed by Genivar Engineering. This review identified difficulties in determining the air volume requirements in this highly mechanized mine and acknowledged the need of third stage analysis. The third stage review was performed by CANMET-MMSL through a contracted study [Hardcastle, Kocsis & Li, 2006]. The Kidd Creek mine is a copper/zinc underground operation located near Timmins, Ontario. The mine started in 1966 with an open pit and has since extended underground through Mines No.1 through No.3 and now Mine D towards 3,000m depth and its 88th primary level, progressively mining steeply dipping massive sulphide deposits. The mine currently produces more than 7,000 tons/day of ore using the blasthole variation of the vertical retreat 63  mining method, combined with cemented backfill while employing a fleet of more than 200 production and support diesel powered equipment with a total rated power of 25,000 kW. To facilitate the movement of development and production workings to greater depths, the mine has undergone various upgrades to its ventilation system which included a change over from a “push-pull” to an exhaust system with the installation of two 2,600 kW (3,500 hp) surface fans, and more recently the addition of two 3,000 kW (4,000 hp) booster fans 1,800 m below the surface [Hortin et al, 2002]. The objective of all these upgrades was to increase the mine’s total ventilation capacity to 1,220 m³/s. Furthermore, due to the added thermal load of mining at depth, the mine has commissioned a 7.5 MW bulk cooling plant on surface to supplement the natural cooling capacity of a “cold stope”, facilitated by Canada’s winter climate. As a result of these upgrades, the mine’s installed fan power within its primary system increased to 13,600 kW.  5.4.1 First Stage Review - Brainstorming Forum This initial review was a macroscopic analysis, which looked at possible restrictions and high frictional/shock pressure losses along the intake and exhaust airways and to the operating characteristics of the main surface fans. The second and third reviews used long-term future production planning data and historic mining activity data that was stored in the mine’s database, respectively, to accurately determine the mine’s intake air volume requirement in an attempt to minimize ventilation redundancy, which in turn would reduce ventilation costs.  5.4.2 Second Stage Review - Future Production Planning Based Analysis Airflow Calculations Based on Future Production Planning – Iteration #1 This analysis was performed in 2005 by Genivar Engineering, which included Level 46.1 through to Level 88, using the 18-month production plan, from January 2006 to June 2007. It was then extrapolated through to the year 2018 using an “airflow to tonnage” ratio, plus an allowance for leakage and additions for ore haulage on the main ramp at the final stages of the mine’s operating life. Throughout, it was also assumed that the mine would eliminate diesel truck haulage of backfill material.  64  In this analysis, three basic minimum ventilation conditions, generically termed in this iteration as “production”, “drilling” and “miscellaneous”, were derived based upon potential diesel powered equipment activity as follows: •  Production: 25.2 m3/s were allocated to a level for development, production, backfilling or rehabilitation activities; this was sufficient for the combined operation of the mine’s standard production LHD and a shotcrete hauler.  •  Drilling: 11.1 m3/s were allocated to a drilling level; this was sufficient for a Cubex drill and a small service vehicle.  •  Miscellaneous: 14.5 m3/s were allocated to a level with minimum planned activity, sufficient for the Toro 1400 LHD.  In this iteration, no individual air volume was allocated to the inactive levels. However, an additional global allowance of 20% was allocated to account for leakage into these inactive areas and old parts of the mine. Based upon these air volume allocations, and without diesel based haulage of backfill material, the minimum required volume was predicted for each month of year 2006, with an average of 585 m3/s, and a highest of 633 m3/s. Furthermore, according to these values and extrapolation through to the year 2018 using an “airflow to tonnage” ratio, the highest ventilation demand was predicted to be a minimum of 916 m3/s for the years 2013 & 2014. These predictions would tend to indicate that the mine should have more than sufficient ventilation delivery capacity and that it should be able to operate more efficiently at a lower volume than currently supplied [Hardcastle, Kocsis & Li, 2006]. Revised Airflow Requirements Based on Operational Logistics – Iteration #2 Due to the disparity between the preceding predictive analysis and traditional ventilation practice, the ventilation design assumptions and minimum requirements were re-evaluated in regard to their operational applicability. This analysis was performed by CANMET-MMSL. In this iteration, one major consideration was the need to avoid uncontrolled recirculation within the auxiliary ventilation system when only one large diesel unit was operating on a level. To avoid recirculation at the fresh air “pickup” location of the auxiliary fan, a minimum airflow velocity of 0.25 m/s needed to be maintained flowing past the auxiliary fan. Based on typical airway dimensions at the Kidd Creek mine, an oversupply of 6.2 m3/s would be required with the following results:  65  •  Increase the “production” air volume allocation from 25.2 m3/s to 29.0 m3/s to accommodate two LHDs. Historical data showed that more that one LHD on a production level was a regular occurrence.  •  Increase the “drilling” air volume allocation from 11.1 m3/s to 14.0 m3/s. In the previous iteration, the air volume allocation for the service vehicle was insufficient in order to avoid recirculation within the auxiliary ventilation system.  •  Increase the “miscellaneous” air volume allocation from 14.5 m3/s to 20.7 m3/s to provide an oversupply at the auxiliary fan and again prevent recirculation.  Furthermore, considering the mining levels are continually switching from active to nonactive states, in order to minimize the potential differential pressures capable to induce leakage in the ventilation system, mining levels with “zero” air volume would be impractical. Here, a useful allowance would be a minimum air volume of 3.5 m3/s, which would be sufficient to dilute the diesel exhaust contaminants generated by the mine’s small service vehicles (<75 hp). Table 9 provides a sample of the airflow allocation from these two iterations plus a summary for Level 46.1 through to Level 88 for the Kidd Creek Mine. The net result in increasing the airflow requirements to minimize leakage plus the minimum flow allocations for the inactive levels was an increase in the predicted average and maximum flow demands by 21-23.5%, to 722 and 765 m3/s. Upon considering the changes in the mine’s annual production tonnage, the maximum air volume required in 2013/2014 now becomes 1,093 m3/s. These new predictions still tend to indicate that the mine should have more than sufficient ventilation delivery capacity and that it should be able to operate at a lower volume than currently supplied Hardcastle, Kocsis & Li, 2006].  66  Table 9: An example of airflow allocations per level (in m3/s) and for all the lower mine based on the 18-month production plan before and after allowances for leakage and to avoid recirculation (shaded areas represent air volume substitutions for production activities, to provide minimum airflow velocities or a to prevent uncontrolled recirculation within the auxiliary system) [Hardcastle, Kocsis & Li, 2006]  67  5.4.3 Third Stage Review - Retrospective Activity Based Analysis based upon the Mine’s Database (SIMS) One of the major failings of the future production planning based analysis (2nd stage review) is that they use production planning data generalized for a specific time span. In this instance the airflow requirements were initially determined on a month by month basis for 2006, and then extrapolated by year through to 2018. This would be sufficient for airflow determination if the ventilation was continually adjusted to match production activity However, in reality, the scheduling of production and redistribution of supporting ventilation is rarely this simple or timely. At Kidd Creek, at the end of every shift, the activity that has taken place is recorded in an electronic database referred to as SIMS. This data details what production/primary diesel equipment was operating, its location(s) and for how long. This equipment data is exportable to standard spreadsheet software where it can be summarized through MSExcelTM spreadsheet based (pivot table) analyses. Consequently, it can be used to generate retrospective air volume requirements based upon engine size and the number of vehicles reporting to each production level. Similar to the preceding analysis, these demands can be qualified to prevent recirculation within the auxiliary ventilation systems depending upon the number and type of diesel vehicles reported. These airflow requirements can then be averaged for various time-spans, which could represent the frequency with which the ventilation is redistributed to the production levels. The following sections consider the airflow requirements assuming the ventilation is redistributed monthly, weekly or daily based upon the day-shift SIMS records of January 2005 through March 2006. Retrospective Activity-Based Analysis Using MS Excel Spreadsheets (Pivot Tables) Table 10 represents a sample of the “monthly” maximum required airflow per level and their summation under two operational conditions, with and without diesel vehicle placed backfill. For the most part, this table shows that on a monthly basis few levels are devoid of any activity, consequently most would require a significant volume. This table also serves to show that unless the ventilation is adjusted regularly it will result in high air volume requirements. With diesel placed backfill and with the mine’s current ventilation practice, the 68  monthly total intake air volume varied between 968 m3/s and 1,559 m3/s. This later volume is well in excess of the mine’s capacity and hence the belief that the ventilation system maybe inadequate [Hardcastle, Kocsis & Li, 2006]. Table 11 provides samples of the “weekly” and “daily” maximum required airflow per level and their summation. Of note within this table is the increasing frequency in the number of days each level only requires the minimum leakage flow, and how this translates into significantly lower overall requirements [Hardcastle, Kocsis & Li, 2006]. Figure 11 and Figure 12 detail the results of the pivot table analyses in graphical terms showing how the total required volume for the lower mine and the number of active levels varied as a function of airflow redistribution frequency. These graphs also include the “future” production planning based predictions derived from long range production planning. Figure 11 shows the benefit of increasing the frequency with which the ventilation is redistributed on the production levels. This graph shows that for the time period analyzed, the average maximum required air volume decreased from 983 m3/s when redistributed “monthly”, to 681 m3/s when redistributed “weekly”, and to 400 m3/s when redistributed “daily”. For comparison, the average monthly production planning based analysis for the period of January - March 2006 was predicted at 632.5 m3/s. Figure 12 shows that by increasing the mine’s ventilation redistribution frequency from monthly to weekly and daily, can significantly reduce the number of active mining levels. Here, the average number of active levels out of a potential 35 is 31.2, 25 and then 17.5. For comparison, the “future” production planning based analysis predicted on average 24 levels as active [Hardcastle, Kocsis & Li, 2006].  69  Table 10: Sample of the lower mine airflow requirements per level (in m3/s) by “month” derived from historical equipment deployment data (shaded areas denote substitutions for minimum airflow or required airflow to prevent recirculation within the auxiliary system) [Hardcastle, Kocsis & Li, 2006] Requirement/ Month 2005-06 January February March Activity Count Average Maximum Minumum Requirement/ Month 2005-06 January February March Activity Count Average Maximum Minumum  46 21.4 6.9 27.9 13 16.4 32.1 3.5  47 25.7 3.5 22.4 12 15.0 25.7 3.5  48 3.5 18.6 18.6 13 19.6 37.9 3.5  49 38.5 61.4 61.0 15 34.2 61.4 16.1  46 21.4 6.9 27.9 13 16.4 32.1 3.5  47 25.7 3.5 18.6 12 14.8 25.7 3.5  48 3.5 18.6 18.6 12 18.1 37.9 3.5  49 3.5 35.9 23.6 13 18.9 35.9 3.5  Backfill Haulage with Diesel Powered Equipment 51 52 53 54 56 57 83 53.0 18.6 58.6 3.5 35.3 20.7 71.8 50.9 14.0 59.9 48.4 36.0 20.7 86.9 50.9 31.6 27.5 18.6 20.7 3.5 69.5 12 11 11 10 12 8 15 27.1 14.5 22.5 18.3 20.7 12.1 40.0 53.0 31.6 59.9 48.4 36.0 23.1 86.9 3.5 3.5 3.5 3.5 3.5 3.5 7.4 Non-diesel based Backfill 51 52 53 54 56 57 83 23.6 18.6 46.2 3.5 35.3 14.7 40.6 20.7 14.0 21.4 22.9 26.9 4.0 46.3 20.7 31.6 14.7 18.6 20.7 24.8 55.0 11 8 11 9 12 8 14 13.8 11.3 15.8 13.7 18.9 11.0 28.7 26.9 31.6 46.2 26.9 35.3 24.8 55.0 3.5 3.5 3.5 3.5 3.5 3.5 3.5  84 84.5 79.1 44.1 14 36.8 84.5 3.5  86 21.1 74.1 87.4 15 48.0 101.7 3.7  87 3.5 20.7 20.7 14 44.6 85.2 3.5  88 3.5 3.5 3.5 8 50.1 122.6 3.5  84 70.0 64.6 29.6 13 30.2 70.0 3.5  86 21.1 59.6 72.9 15 40.2 72.9 3.7  87 3.5 3.5 3.5 11 31.0 70.7 3.5  88 3.5 3.5 3.5 8 37.0 85.3 3.5  Grand Total 1314.8 1477.5 1559.3 max 15 1260.6 1559.3 968.2 Grand Total 1005.3 1088.4 1125.2 max 15 983.4 1163.6 807.6  70  Table 11: Sample of the lower mine airflow requirements per level (in m3/s) by “week” and “day” derived from historical equipment deployment data (shaded areas denote substitutions for minimum airflow or required airflow to prevent recirculation within the auxiliary system) [Hardcastle, Kocsis & Li, 2006] Requirement/ Week 2005-06 Week 1 Week 2 Week 3 Week 4 Activity Count Average Maximum Minumum Requirement/ Day 2005-06 08-Jan-05 09-Jan-05 10-Jan-05 11-Jan-05 12-Jan-05 13-Jan-05 14-Jan-05 Activity Count Average Maximum Minumum  46 21.4 3.5 17.4 14.7 30 9.1 32.1 3.5  47 3.7 11.8 18.6 25.7 33 9.0 25.7 3.5  48 3.5 3.5 3.5 3.5 32 9.7 37.9 3.5  49 3.5 3.5 3.5 3.5 46 13.1 35.9 3.5  51 3.7 20.7 20.7 23.6 26 8.2 26.9 3.5  52 17.4 18.6 3.5 17.4 21 7.5 31.6 3.5  46 3.5 3.5 3.5 3.5 3.5 3.5 3.5 67 5.0 32.1 3.5  47 3.7 3.7 3.7 11.8 3.7 3.5 14.0 106 5.7 25.7 3.5  48 3.5 3.5 3.5 3.5 3.5 3.5 3.5 110 5.6 37.9 3.5  49 3.5 3.5 3.5 3.5 3.5 3.5 3.5 209 8.5 35.9 3.5  51 3.7 3.7 3.7 3.5 3.7 20.7 3.5 96 6.0 26.9 3.5  52 3.5 3.5 3.5 3.5 3.5 3.5 3.5 58 4.8 31.6 3.5  Non-diesel based Backfill 53 54 56 57 37.2 3.5 3.5 14.7 24.8 3.5 7.4 14.7 27.3 3.5 35.3 3.5 46.2 3.5 3.7 3.5 24 35 41 19 9.1 11.0 12.8 7.0 46.2 26.9 35.3 23.1 3.5 3.5 3.5 3.5 Non-diesel based Backfill 53 54 56 57 37.2 3.5 3.5 14.7 24.8 3.5 3.5 14.7 24.8 3.5 3.5 3.5 23.6 3.5 3.5 3.5 16.1 3.5 3.5 3.5 18.6 3.5 3.7 6.9 19.6 3.5 7.4 3.5 97 135 143 51 6.3 7.2 7.1 4.6 46.2 26.9 35.3 23.1 3.5 3.5 3.5 3.5  83 36.8 40.6 32.1 36.8 50 19.7 55.0 3.5  84 70.0 59.6 52.0 66.5 37 18.4 70.0 3.5  86 3.5 3.5 3.5 21.1 59 27.1 72.9 3.5  87 3.5 3.5 3.5 3.5 45 23.6 70.7 3.5  88 3.5 3.5 3.5 3.5 35 31.8 85.3 3.5  Grand Total 693.6 674.6 709.8 742.9 max 65 681.4 867.3 438.8  83 26.1 40.6 19.2 34.4 3.5 6.9 35.6 297 10.9 55.0 3.5  84 46.6 18.7 52.0 24.5 59.6 15.2 26.1 165 9.8 70.0 3.5  86 3.5 3.5 3.5 3.5 3.5 3.5 3.5 352 14.7 72.9 3.5  87 3.5 3.5 3.5 3.5 3.5 3.5 3.5 301 15.1 70.7 3.5  88 3.5 3.5 3.5 3.5 3.5 3.5 3.5 235 21.7 85.3 3.5  419.3 409.9 373.3 388.3 378.4 395.1 425.4 max 455 399.7 574.6 126.0  71  Figure 11: Comparison of the lower mine’s ventilation requirements derived from historic production data (daily, weekly, monthly) and from future monthly production plans [Hardcastle, Kocsis & Li, 2006]  72  Figure 12: Comparison of the number of active levels within the lower mine derived from historic production data (daily, weekly, monthly) and from future monthly production plans [Hardcastle, Kocsis & Li, 2006]  73  SIMS Data Analysis Limitations One of the limitations of the available SIMS data is that the mining equipment does not have a “time-stamp”. Consequently, when the time reported is less than the full shift length, or when a vehicle is reporting to more than one level, it would be extremely difficult to establish when a vehicle was working in a specific area. This could cause the pivot table analysis to overestimate the air volume requirements, as two or more vehicles reporting to the same level may be there at different times. For example, the MSExcelTM spreadsheet based (pivot table) analyses, produced airflow requirements for more than 100 m3/s on several production levels. These flow requirements were considered very high when compared to the 30 m3/s, which is currently being allocated. These high air volume requirements were typically generated on levels where multiple vehicles, up to 5, were recorded as “active”. Although recognized as overestimated, due to limited information with respect to where these vehicles were operating, these air volumes could not be totally discounted when determining the mine’s intake air volume.  5.4.4 Variance between “Future Production Planning” and “Retrospective Activity Based” Methods of Analysis The results from the various analyses as summarized in Figure 11, show significant differences between the airflow requirements determined for the lower part of Kidd Creek mine as derived from the historical SIMS activity data and the 18-month production plan. The differences are a result of the following: •  The initial production planning analysis (2nd Stage Review – Iteration #1), using the 18-month plan did not account for the additional requirements of an auxiliary ventilation system to avoid recirculation on a level where only one vehicle was operating. This analysis also assumed that temporarily inactive levels would not be ventilated, whereas in reality they would still be ventilated such that minimum airflow velocities are maintained along the main airways.  •  The amended production planning analysis (2nd Stage Review – Iteration #2), although taking account of the auxiliary fan requirements and a minimum flow per non-active level, was still based upon the 18-month production plan. This is an “ideal” plan that is more focused on ensuring tonnage than the deployment of such resources as ventilation. 74  •  Both future production planning analyses (Iteration #1 & Iteration #2) were based upon averaged production activity and do not include any maximum concentrated activity demands.  •  The retrospective mining activity analysis using SIMS data (3rd Stage Review), was based upon the “observed” discontinuous production activity details, which highlights maximum demands as opposed to the idealized average activity.  Despite the review, which resulted in additional air volume allocations, the monthly production based analyses still appear to underestimate the overall flow requirement when compared to the historical equipment activity analysis. The primary reason for this is due to the oversimplification of the future production plans. On the contrary, the historical based analyses seem to overstate the overall flow requirement on a monthly basis due to the potential multiple accounting of diesel vehicles that report to more than one location per shift. However, regardless, of the actual flow predictions, the historical based analyses will give a truer estimation of the number of levels that could be active; this value is independent of the number of vehicles reporting to each level. Both analyses show that for the mine to operate efficiently, it must regularly redistribute its ventilation. However, to achieve a flow requirement comparable to the production plan based requirement, as shown by the historical analysis, the mine has to consider at least redistributing its flow weekly.  5.4.5 Ventilation Redundancy at the Kidd Creek Mine - Discussions Determining the current overall efficiency of the ventilation system at the Kidd Creek Mine could be subjective. The reductions between monthly, weekly and daily, as shown in Figure 11, despite being considerable, will not be the same when it comes to implementing an increased airflow management and control infrastructure. This is because the current ventilation system is incapable of meeting the maximum predicted airflow requirements that would allow the mine to achieve the future planned production rates with the ventilation being redistributed monthly, as was the mine’s standard practice. Currently, the mine appears to have sufficient ventilation capacity if its airflow would be redistributed somewhere between weekly and monthly. Therefore, the only major improvement that would minimize redundancy and thereby improve the system’s relative 75  efficiency is to redistribute the airflow for the production levels more frequently than weekly. This in all likelihood would necessitate some form of ventilation automation. In the historical analyses, summarized in Figure 11, it can be seen that a more frequent redistribution, from weekly to daily could reduce the average demand by 40% and peak by 26%. Due to the magnitude of the mine’s current electrical usage, this order of reduction provides the justification to consider some level of automation as the mine continues to develop its lower levels. Minimizing ventilation costs in large mechanized mines is not straightforward due to the dynamic and variable nature of production process constantly redefining where ventilation is required. Using future production planning data can greatly underestimate the required flow unless some consideration is given to how that ventilation is delivered to the point of application. In this analysis, the need to prevent uncontrolled recirculation within auxiliary ventilated locations and allocations to non-productive levels significantly increased the volumetric predictions of the future production based analysis. However, even with these considerations accounted for, the predictions could still underestimate the ventilation demand due to their oversimplification of where production was taking place, and/or the assumption that the ventilation would be adjusted accordingly. The use of historical data can capture the ever changing and dynamic nature of mining in a more detail, but it can still be inadequate. For example, at Kidd Creek, within a month, the future production based analysis (2nd stage review) only showed on average 24 mining levels as being active, whereas the historical data showed 31 levels requiring ventilation for production related activities However, the historical equipment data analyses (3rd stage review) seemed to overestimate the actual volumes required. This was due to the recorded diesel activity data not having any “time-stamp” by which to correct the air volumes for multiple vehicles reporting to a particular level but at different times. Despite this deficiency, the major benefit of the historical equipment data (SIMS) analysis is its ability to show: •  The significant benefits of increasing the adjustment and redistribution frequency of the ventilation system, and  •  The importance of being able to set the mine’s overall flow appropriately when one increases the frequency of redistributing the airflow. 76  5.5  Challenges and Difficulties in Determining Ventilation Redundancy in Large and Deep Multilevel Operations  These two ventilation studies have demonstrated the difficulty, complexity and labour intensiveness in determining potential ventilation redundancy in two large and deep multilevel metal mines by means of conventional mine ventilation analyses techniques. In the first case study “Determination of Ventilation Redundancy at Creighton Mine – Vale Inco”[Hardcastle & Kocsis, 2004], the mine activity logs have shown that maximum primary ventilation is required 56% of the time in a standard working week and for the remaining 44% of the time it could be reduced to a lower rate. However, determining the ventilation redundancy within the mine’s auxiliary ventilation system mine-wide was problematic due to a large number of auxiliary fans (>200 in single/series configuration) and the lack of data as to when any of these fans were operating. The mine activity logs have also shown that on average an auxiliary fan in the production area may be required as little as 20% of the time. In this study it was also difficult to predict how the power requirements of the auxiliary system will change with each stage of the mine’s development as it would be unrealistic to assume that they would remain constant especially with the addition of more and possibly larger auxiliary fans. Furthermore, during 125 days of mining activity monitoring, the temperature, humidity and gas monitoring units generated an extremely large amount of electronic data. As a result, significant resources were required in order to download and process these data in graphs that were able to show when activities started and ceased as well as the underground climatic conditions they created. The second case study: “Ventilation Utilization Efficiency of the Kidd Creek Mine – Xstrata PLC” [Hardcastle, Kocsis & Li, 2007), showed again that determining ventilation redundancy in a large and mechanized metal mine is not straightforward due to the dynamic nature of production processes continually redefining where and how much ventilation is required to maintain safe environmental conditions for the underground workforce. In this study, to determine potential ventilation redundancy and improve the efficiency of the mine’s ventilation systems a three stage analysis was performed.  77  The study showed that the future production based analysis (2nd stage review) can greatly underestimate the mine’s intake air volume unless some consideration is given to prevent uncontrolled recirculation and provide minimum air volumes to the non-productive levels. Even with these considerations in place, the future production based analysis predictions could still underestimate the ventilation requirements due to insufficient information and oversimplification with respect to where production was taking place, in other words which stopes were active. The use of SIMS data analysis (3rd stage review) can capture the dynamic nature of mining activities. However, the SIMS data analysis seemed to overestimate the mine’s air volume requirement due to the recorded diesel equipment activity data not having a “time-stamp” by which to correct for multiple production vehicles reporting to a production level but at different times.  5.6  The Need to Change Current Ventilation Design Practices  Currently, the underground ventilation systems are designed towards the worst-case-scenario with respect to ventilation demand, which usually occurs within a future time period of the mines’ operating life. This means that during the early stages of their life these mines can be significantly over-ventilated. Such ventilation systems are inefficient and wasteful and this design approach must change if the Canadian mining operations are to remain competitive. This ventilation design approach based upon the assumption of peak production and delivery of maximum air volumes to all potential active and inactive production areas was in the past and even recently challenged and criticized [Hardcastle et al, 2004] [Allen et al, 2008] [O’Connor, 2008]. Other authors expressed their concern regarding the current ventilation design criterion according to which the mines’ intake air volumes are determined based entirely on the total diesel engine fleet capacity and a statutory requirement per kW of rated engine power [Brake et al, 2008]. This is because the current methods of airflow evaluation do not have the capability to address and take into account the dynamic nature of the development and production operations that are constantly redefining where and how much ventilation is needed at a certain point in time of the mines’ life. As a result, such airflow evaluation methods will inevitably result in either overestimating or underestimating the mines’ airflow requirements.  78  When designing a new ventilation system or optimizing existing ventilation systems, in order to minimize ventilation redundancy within the production areas and throughout the mines, the following questions need to be asked [Hardcastle & Kocsis, 2002]: •  Are all production areas in the mine truly active? Are there any potential production workings, which due to planning or other operational reasons became temporarily inactive?  •  Do all these production areas need the same amount of airflow? Would it be possible to reduce the air volume within the temporarily inactive areas such that only minimum airflow velocities are maintained along the main levels?  •  During the mining cycles, do these production workings require a constant air volume, in other words does the explosive loading operation performed using a small electric loader (e.g. 40 kW) require the same amount of air as the mucking operation performed with a large diesel LHD (e.g. 250 kW)?  •  In a production stope with multiple draw-points, is it possible to manage the auxiliary airflow delivery such that fresh air is delivered only to the drawpoint(s) where the fragmented ore is loaded and removed by LHDs?  •  For how long should ventilation be supplied at any location? Within an active production level would it be possible to reduce the air volume during scheduled activity breaks (e.g. lunch break)?  •  For what period of the production shift is the entire mine active? Is there a consistent pattern during which the entire mine becomes inactive (i.e. shift changes), during which both the primary and auxiliary ventilation delivery can be scaled down? In some instances airflow delivery would be required continuously, such as to control blast fume clearance. However, is this airflow requirement as demanding as for diesel usage?  •  In deep mines, ventilation is required to remove heat, but considering that strata is an infinite source of heat, what would be the result if the airflow delivery was stopped for a few hours or over a longer time period (e.g. weekends). Obviously, the mining areas would heat up, but how long would it take for the mining area to return to adequate climatic conditions once the airflow delivery is resumed?  79  •  Actually, in highly mechanized metal mines, where large auxiliary fans are employed, they can significantly increase the temperature of the air within the ducting system downstream (e.g. ∆TDB = 2-3 0C).  •  Have the original ventilation system design criteria been exceeded?  When answers to these questions or even some of these questions are given and taken into account, significant savings in the mines’ intake air volumes can be achieved. Furthermore, due to the cubic relationship between the mine’s supplied air power and the airflow (Power α Flow3), even small reductions in the mines’ intake air volume can generate significant savings in underground energy consumption and operating costs. Fortunately, all the above mentioned questions can be imported and their effects evaluated through discrete-event mining process simulations performed on mine models that integrate production, development and secondary operations. With increasing pressures to reduce underground energy consumption and the carbon footprint of the Canadian mining industry, it is becoming increasingly important to develop a new ventilation design concept and show the mining industry how discrete-event process simulations performed on a mine model that integrates production, development and secondary operations can be used to develop an activity based “life-cycle” airflow demand schedule using AutoModTM simulator. This activity based “life-cycle” airflow demand schedule as opposed to maximum ventilation demand would then become the basis for the design of a mine’s primary and auxiliary ventilation infrastructure. Furthermore, the integration of mining process simulation and ventilation simulation could also allow the exploration of options to negate, and possible reverse the ventilation drivers (e.g. increasing depth, mechanization) that under traditional ventilation practice would dictate increased intake air volumes and energy consumption. With this ventilation design approach, the mining industry could reduce its energy consumption and consequently its carbon footprint directly through reduction in heating fuel usage and more significantly through reduced underground electrical usage.  80  6.  MINING PROCESS MODEL TM AutoMod SIMULATOR  DEVELOPMENT  USING  Process simulation can be defined as the imitation of a real-world process or system over a period of time. Simulation usually involves the generation of an artificial history of the system (i.e. model development), which when extensively analyzed generates inferences concerning the operating characteristics of the real system. Today, simulation has become an indispensable problem solving methodology for the solution of many “real-world” problems. Simulation and animation can be used to describe and analyze the behaviour of a system, ask “what-if” questions about the real system and simulate scenarios generated by these questions in order to support the design of real systems. Both existing and conceptual systems can be simulated by means of computer modelling techniques.  6.1  Development of the “Mining Process” Model  To determine a mine’s activity based air volume requirements, a generic “Mining Process Model” was developed in collaboration with Penguin Automated Systems Inc, under a contract agreement with CANMET-MMSL, using AutoModTM. The initial version of the “Mining Process” model with its static and process systems was coded by Penguin Automated Systems Inc. The process logic was later modified by the author, and throughout its subsequent versions additional subroutines were added to the main process systems. These enabled the “Mining Process” model to simulate mining and fan activities in real-time, based upon planned production operations schedules entered by the user.  6.1.1 Mining Processes and Systems The static system of the “Mining Process” model includes two production stopes that employ the vertical retreat mining method (each stope with a drilling and a mucking level), two developments and two ore and waste passes associated to the production and development workings, respectively. The drilling and mucking levels of the production stopes (VRM 1 & VRM 2), the development headings as well as the ore & waste passes have their own independent auxiliary ventilation system. Within both production stopes, the drilling level consists of an overcut that accommodates eight possible drilling locations. The drill pattern used is similar to the longhole pattern of the sublevel stoping method, however in this case, the holes are loaded and blasted from the sublevel at the top of the stope in horizontal slices. 81  This requires that the drillholes first be sealed with plugs that can be installed from the above sublevel. The 205 mm diameter drillholes are then loaded with a fixed height of charge and blasted into the undercut. This operation is repeated until the blasting horizon approaches the overcut of the stope. In this model, the required fresh air throughout each drilling level of both VRM stopes is supplied through a 38” auxiliary steel duct with the assistance of one auxiliary fan. The haulage level of each production stope consists of a 5.2m x 4.8m drift that connects the working area to the main access ramp and from here to the ore-pass. The working area of each production stope includes 8 draw-points (in a 2 x 4 layout), with four draw-points located on one side and 4 draw-pints located on the opposite side of the mining block’s centre line. From the draw-points the blasted ore is transported along the haulage drifts and the main access ramp to two ore-passes (Ore-pass #1 & Ore-pass #2) using two ST8 Wagner LHDs, 242 kW (325 hp) each. The haulage drifts of both production stopes have their own independent auxiliary ventilation system. The fresh air to the draw-points is delivered along 48” & 38” steel ducts with the assistance of two auxiliary fans. Within the production stopes the ducts and associated auxiliary fans that deliver fresh air to draw-points 1-5, 1-6, 1-7, 1-8 (stope #1) and 2-5, 2-6, 2-7, 2-8 (stope #2), have been termed as “scavengers”. This is because these fans are pulling fresh air from another independent system, in this case the ducting system through which fresh air is delivered to draw-points 11, 1-2, 1-3, 1-4 (stope #1) and 2-1, 2-2, 2-3, 2-4, 2-5 (stope #2), respectively. Therefore, depending upon the drawpoint from where the fragmented ore is pulled, one or two of the auxiliary fans associated to each haulage level could be operational (see Figure 13). The developments (i.e. Development #1 & Development #2) are located below the stopes, both driven off of the main access ramp. The fragmented waste resulted from a blasted development round is transported along the two haulage drifts and the main access ramp to the waste passes (i.e. Waste-pass #1 & Waste-pass #2). The development headings and both waste passes have their own independent auxiliary ducting system (see Figure 13).  82  Figure 13: Representation of a mining block with 2 production stopes (VRM 1 &VRM 2), 2 developments, 2 ore & rock passes  Drilling level  Main Access Ramp VRM1 mucking level  Draw-points  Drilling level Ore-pass 1  VRM 1 Draw-points: 1-5 1-6 1-7 1-8 Draw-points: 1-1 1-2 1-3 1-4  Parking bay 1  VRM 2 VRM2 mucking level  Draw-points  Draw-points: 2-1 2-2 2-3 2-4  Ore-pass 2 Parking bay 2 Development 1  Waste-pass 1 Development 2  Waste-pass 2  Draw-points: 2-5 2-6 2-7 2-8  83  The “Mining Process” model has the following main components: •  One “process system”, which includes 320 “subroutines”, saved in 21 independent source files. These subroutines create the process logic that controls the mining activities within the production stopes, developments, along the haulage drifts and the ore and waste passes. These subroutines also contain the logic that controls the activity of all auxiliary fans assigned to the mining block which are turned “On” as soon as the mining equipment enters the working areas and “Off” with the completion of mining activities or departure of the mining equipment from the working areas.  •  Five “path moving” systems, which graphically represent the drilling and mucking levels of the production stopes, developments, ore/rock passes, haulage drifts and the main ramp system (see Figure 14). The path moving systems also include a “work list” assigned to the longhole drill, the Jumbo-drill, explosive loader and LHDs, which in conjunction with the process logic control the deployment and operation of the mining block’s production and development fleet.  The process logic of the simulation model was developed according to the ore extraction sequence, longhole drilling and development cycles of a typical mining block that employs the vertical retreat mining method. The activity logic incorporates all fan activity subroutines as well, coded such that an appropriate amount of fresh air is delivered to the production and development workings based upon the mining activities and consequently the type and size of the mining equipment. In other words, within the mining blocks, fresh air to the production workings is delivered when, where and only as long as the mining equipment operates. In addition the subroutines of the main process logic include the following operational conditions: •  Within the stopes, during the same time period, the LHDs are not allowed to pull fragmented ore from a draw-point located below an active drilling area of the overcut.  •  Within the mining block, on the drilling and mucking levels individual activities can occur concurrently. However, mucking can only follow explosive loading and blasting activities. Furthermore, the blasted ore from a draw-point must be completely removed before the longhole drilling operation above that particular drawpoint resumes.  84  Figure 14: Locations of the auxiliary fans in this multi-level mining block – “Mining Process” model Total Intake Air Volume to the Mining Block Scav1_drill  Main Access Ramp Scav_fan1_scoop  VRM Stope 1 Draw-points: 1-5 1-6 1-7 1-8  Fan1_scoop Orepass1 Scav2_drill  Draw-points: 1-1 1-2 1-3 1-4  Ore-pass 1 Scav_fan2_scoop Fan2_scoop Orepass2  Draw-points: 2-5 2-6 2-7 2-8 Draw-points: 2-1 2-2 2-3 2-4  Dev1_fan_scoop  Ore-pass 2  Waste-pass 1  Dev1_fan  Waste-pass 2 Dev2_fan  Development 1  Dev2_fan_scoop  Development 2  85  6.1.2 Ventilation System The auxiliary ventilation system of the “Mining Process” model consists of 10 independent delivery systems. The fresh air to the production and development workings as well as to the ore and waste passes is delivered along 48” & 38” steel ducts with the assistance of 12 auxiliary fans. The return air is exhausted to surface along the haulage drifts, then into the mine’s return air system. The auxiliary fans assigned to the mining block are located inside the steel ducts through which the required air volumes are delivered to the working areas. The locations of the auxiliary fans assigned to the mining block are shown in Figure 14, as follows: •  Drilling level of the VRM #1 stope: with one auxiliary fan, namely “Scav1_drill” located inside a 38” steel duct.  •  Mucking level of the VRM #1 stope: with 2 auxiliary fans, “Fan1_scoop” serving draw-points 1-1, 1-2, 1-3, 1-4 and “Scav_fan1_scoop” serving draw-points 1-5, 1-6, 1-7, 1-8, located inside 48” and 38” steel ducts, respectively.  •  Drilling level of the VRM #2 stope: with one auxiliary fan, namely “Scav2_drill” located inside a 38” steel duct.  •  Mucking level of the VRM #2 stope: with 2 auxiliary fans, “Fan2_scoop” serving draw-points 2-1, 2-2, 2-3, 2-4 and “Scav_fan2_scoop” serving draw-points 2-5, 2-6, 2-7, 2-8 located inside 48” and 38” steel ducts, respectively.  •  Haulage drift connecting the main ramp to Ore-pass #1: with one auxiliary fan, “Orepass1” located inside a 48” steel duct.  •  Haulage drift connecting the main ramp to Ore-pass #2: with one auxiliary fan, “Orepass2” located inside the 48” steel duct.  •  Development #1 drift: with one auxiliary fan “Dev1_fan_scoop” located inside a 38” steel duct.  •  Development #2 drift: with one auxiliary fan “Dev2_fan_scoop” located inside a 38” steel duct.  •  Haulage drift connecting the main ramp to Waste-pass #1: with one auxiliary fan, “Dev1_fan”, located inside a 38” steel duct  •  Haulage drift connecting the main ramp to Waste-pass #2: with one auxiliary fan, “Dev2_fan”, located inside the 38” steel duct 86  The minimum air volume requirements along the haulage drifts during lunch breaks, shift changes and during time periods when no mining equipment operates were determined according to minimum airflow velocity considerations as follows: •  Minimum air volume along the stope haulage levels: QHmin = 7 m3/s  •  Minimum air volume along the development haulage levels QDmin = 7 m3/s  •  Minimum air volume along any section of the of the access ramp: QRmin = 14 m3/s  During the scheduled production down times, shift changes and other inactive time periods, to deliver the minimum air volumes along the production and development haulage drifts, the main access ramp and to avoid airflow recirculation at the auxiliary fan “flow pick-up” locations, the combined minimum air volume required by this multi-level mining block is: QMin-ModelB = 33.3 m3/s. During the mining operations, the minimum air volumes required at the production drawpoints, development headings and along the haulage levels of the ore & waste passes were determined according to the diesel exhaust dilution criterion of 0.063 m3/s/kW (100 cfm/bhp) of rated diesel power. Furthermore, the required air volumes were determined based upon the maximum size and number of diesel units that might operate at any time within the development and production workings of a typical multi-level mining block that employs the vertical retreat mining method, as follows: •  For the drilling levels of the VRM stopes: − One Cubex 6200D-20ITH longhole drill (125 hp)  = 6.0 m3/s  − One man carrier (125hp)  = 6.0 m3/s  Total for VRM drilling level •  For the mucking levels of the VRM stopes: − One ST8 Wagner LHD (325hp)  = 15.3 m3/s  − One man carrier (125hp)  = 6.0 m3/s  Total for VRM mucking level •  = 12.0 m3/s  = 21.3 m3/s  Along each haulage drift to Ore-passes #1 & Ore-pass #2: − One ST8 Wagner LHD (325hp)  = 15.3 m3/s  − One man carrier (125hp)  = 6.0 m3/s  Total for each haulage drift to OP #1 & OP #2  = 21.3 m3/s  87  •  For each development level: − One ST3.5 LHD (185hp)  = 8.8 m3/s  − One Jumbo-Drill (125hp)  = 6.0 m3/s  Total for each development heading •  Along the haulage drift to Waste-pass #1: − One ST3.5 LHD (185hp)  = 8.8 m3/s  − One man carrier (125hp)  = 6.0 m3/s  Total for haulage drift to Waste-pass #1 •  = 14.8 m3/s  = 14.8 m3/s  Along the haulage drift to Waste-pass #2: − One ST3.5 LHD (185hp)  = 8.8 m3/s  − One man carrier (125hp)  = 6.0 m3/s  − Avoid airflow recirculation at aux. fan  = 3.7 m3/s  Total for haulage drift to Waste-pass #2  = 18.5 m3/s  Based upon these flow requirements the air volume that needs to be delivered by each auxiliary fan assigned to this mining block are: •  VRM #1 drilling level fan: Q(Scav1-drill) = 12.0 m3/s  •  VRM #1 mucking level fans: − Q(Fan1-scoop) = 21.3 m3/s − Q(Scav-fan1-scoop) = 18.0 m3/s  •  VRM #2 drilling level fan: Q(Scav2-drill) = 12.0 m3/s  •  VRM #2 mucking level fans: − Q(Fan2-scoop) = 21.3 m3/s − Q(Scav-fan2-scoop) = 18.0 m3/s  •  Ore-pass #1 haulage drift fan: Q(Orepass #1) = 21.3 m3/s  •  Ore-pass #2 haulage drift fan: Q(Orepass #2) = 21.3 m3/s  •  Development #1 haulage drift fan: Q(Dev1-fan-scoop) = 14.8 m3/s  •  Development #2 haulage drift fan: Q(Dev2-fan-scoop) = 14.8 m3/s  •  Waste-pass #1 haulage drift fan: Q(Dev1-fan) = 14.8 m3/s  •  Waste-pass #2 haulage drift fan: Q(Dev2-fan) = 18.5 m3/s  88  6.1.3 Input Data The mining process and equipment data such as the time occurrence and duration of production and development operations, namely drilling, explosive loading, blasting, ore/waste haulage, as well as the mining equipment capacities plus the fragmented ore/rock tonnage can be entered in the “Mining Process” model by using an MSExcelTM spreadsheet utility (see Figure 15 & Figure 16). A visual basic application (VBA) then links this spreadsheet utility to the process logic of the model, and then downloads the entered mining activity data into a data file, namely “Data.d” located in the same directory where the “Mining Process” model and its components reside. Before a mining process simulation is performed, the model initiation functions developed within the process logic will upload the mining activity data into the animation module of AutoModTM simulator. The mining activity, equipment data and fan operating duties are then transferred to the various subprocesses of the main process logic that controls all mining operations and auxiliary fan activities within the multi-level mining block. In addition to the previous model, through the use of this data entry spreadsheet utility, the scheduled production delays such as lunch breaks, VRM blasting time periods, shift changes and other temporary down-times can also be entered in the “Mining Process” model (see Figure 15). Furthermore, the logic of the “production delay” subroutine was coded such that the local production down-times that were entered through the spreadsheet utility would only affect a very restricted area of the mining block and its associated activities such as one or both stopes, one or both developments or the ore/waste passes. However, the global type down-times such as VRM blasting time periods (e.g. between the night shifts) would stop any activity within the mining block and dispatch all mining equipment to their associated parking bays. The auxiliary fan operating duties and other ventilation parameters (as defined in section 6.2.2) were entered in the “Mining Process” model using a MSExcelTM worksheet utility, namely the “Fan Volume” worksheet, developed within the “Results” data file, that in turn was generated through mining process simulation (see Figure 17). Furthermore, a visual basic application, namely the “fan Summary” macro would then assign the required air volumes to each individual auxiliary fan and determine the activity-based air volume required by the mining block. 89  Figure 15: Spreadsheet based mine activity input data utility for the “Mining Process” model (i.e. Activity Breaks, VRM1)  90  Figure 16: Spreadsheet based mine activity input data utility for the “Mining Process” model (Development 1, Development 2)  91  Figure 17: Fan volumes data entry using MSExcelTM worksheet  92  This method of fan data entry into the “Mining Process” model presents the following characteristics: •  Unless modifications to the process logic or to the path moving systems are performed, the “Mining Process” model does not need to be saved and re-compiled after the fan operating duties are entered. As a result, significant process simulation time and model compiling problems can be eliminated.  •  During discrete-event process simulation, the “Fan Summary” macro that assigns the required air volumes to the auxiliary fans has the ability calculate the combined activity based air volume that was “handled” by the auxiliary fans as well as the combined activity based air volume that is “required” by the mining block.  •  Under certain operating conditions (i.e. ore/waste haulage), to ensure that adequate climatic conditions are maintained along the production and development haulage drifts and within the production draw-points, it may be necessary that some of the auxiliary fans responsible for delivering fresh air to these locations would need to be turned “On” even before the mining equipment enters the working area. In this respect, the “Fan Volume” worksheet that assigns the required air volumes to the fans has the ability to allocate these pre-programmed time periods to the targeted fans (selected by the user) and take into account these time-periods when calculating the activity based air volume required by the mining bock (see Figure 17).  6.1.4 Output Data Manipulation and Analysis Mining process simulation trials performed on an early version of the “Mining Process” model identified the following model development deficiencies: •  The operating duties of the auxiliary fans assigned to the mining block were entered through coding directly in the process system. Consequently, any changes with respect to the operating duties of the fans were required to be performed by modifying the process logic code. A significant disadvantage of this fan data entry method was that the user of the model was required to be familiar with the structure of the process logic and AutoModTM coding syntax.  •  The mining activity data input spreadsheet utility and the “Mining Process” model initiation functions did not have the capability to take into account scheduled 93  production down-times (e.g. lunch breaks) or unscheduled production delays (e.g. equipment failure) when determining the activity based air volume requirements To eliminate the above mentioned deficiencies, the structure of the process logic that controls all mining and fan activities within the “Mining Process” model was entirely re-designed. The process logic was modulated into a number of 320 sub-processes contained within 21 independent source files. With this development approach, the flexibility of the “Mining Process” model and its ability to generate output data through discrete-event process simulation significantly improved. For example, mining and fan activities within a very restricted area of the mining block (i.e. drawpoint 1-1 of VRM #1) are now controlled by smaller and distinctive subroutines. During discrete-event process simulation these subroutines are always interconnected within the main process logic. The main process logic also contains dedicated model initiation functions that interconnect these subroutines with the mining activity data entry utility. In the “Mining Process” model, the ventilation variables that were defined within the various subroutines have the ability to show when and for how long each particular fan assigned to this mining block was “On” in terms of absolute clock (AC) time. With the completion of a discrete-event process simulation, these fan operating characteristics are transferred from the “Results” data file generated by AutoModTM into the “Custom Report” worksheet (see Figure 18). It should also be mentioned that the beginning of a discrete-event mining process simulation is usually set to correspond to “AC” time zero. A visual basic application, namely the “Fan Summary” macro (see Figure 19), will then divide the total process simulation time-frame (i.e. for this model run 250 hours) displayed in the “Custom Report” worksheet, into very small time intervals that can range from one minute to 60 minutes. Based upon the scheduled production activities provided by the user, the “Fan Summary” macro will determine and display the auxiliary fans that were “On” within a certain time interval (i.e. 5 minutes). This time interval can also be selected by the user. Furthermore, the “Fan Summary” macro will then associate the required air volumes (determined in section 6.2.2 and entered in the simulation model through the “Fan Volume” worksheet utility), to the auxiliary fans and calculate both, the combined activity based air volume that was “handled” by the auxiliary fans as well as the combined activity based air volume that is “required” by the mining block versus the combined air volume that would be required by traditional ventilation. 94  Figure 18: Custom report showing when and for how long each auxiliary fan is operating in hours (absolute clock)  95  Figure 19: “Fan Summary” macro developed for the “Mining Process” model  Through the use of the ”Data Input” spreadsheet utility, various production and development schedules combined with stope extraction and sequencing scenarios can now be entered and evaluated through discrete-event process simulations performed on the “Mining Process” model. For example, a mining process schedule prepared for a typical mining block that employs VRM includes the following activities and their durations: 1) VRM Stope No. 1: •  Longhole drilling using Cubex 6200D-20ITH: 20 hours  •  Explosive loading into longholes: 6 hours  •  Production blasting (horizontal slice) and blasting fumes clearance: 0.5 hours  •  Capacity of the LHD (ST8 Wagner – 325hp) moving ore to the ore-pass #1: 8yd3  •  Blasted ore tonnage: 1,600 tons/blast (200 tons from each drawpoint)  2) VRM Stope No. 2: •  Longhole drilling using Cubex 6200D-20ITH: 30 hours  •  Explosive loading into longholes: 6 hours  •  Production blasting (horizontal slice) and blasting fumes clearance: 0.5 hours  •  Capacity of the LHD (ST8 Wagner) moving ore to the ore-pass: 8yd3  •  Blasted ore tonnage: 2,400 tons/blast (300 tons from each drawpoint)  3) Development No. 1: •  Development face drilling with Jumbo-Drill: 5 hours  •  Explosive loading: 1.5 hours 96  •  Development blasting and fumes clearance: 0.5 hours  •  Capacity of LHD (ST3.5 Wagner) moving waste to the waste-pass: 3.5 yd3  •  Blasted waste tonnage: 200 tons/round  4) Development No. 2: •  Development face drilling with Jumbo-Drill: 6 hours  •  Explosive loading: 1.8 hours  •  Development blasting and fumes clearance: 0.5 hours  •  Capacity of LHD (ST3.5 Wagner) moving waste to the waste-pass: 3.5 yd3  •  Blasted waste tonnage: 250 tons/round  For the simulation model, the starting times as well as the durations of production and development operations in terms of absolute clock (AC), that are relative to the initiation of a discrete-event mining process simulation (ACStart-of-Simulation = 0), were entered using the “Data Input” spreadsheet utility as shown in Figure 20, Figure 21 & Figure 22, respectively. Based upon this operating scenario, the fan activity data generated by AutoModTM simulations performed on the “Mining Process” model is shown in the “Custom Report” worksheet of the “Output Data” file (see Figure 23 & Figure 24). This worksheet shows that each auxiliary fan operating within the mining block has two associated columns. The first column shows the time (in minutes AC) when each auxiliary fan assigned to the mining block was required to be turned “On”. The second column shows the time (in minutes AC) when each auxiliary fan was turned “Off” as the mining operation was completed or when the mining equipment temporarily or permanently exited the development and production working areas. Furthermore, within each row of the “Custom Report” worksheet, the difference between the second and first column shows the time-period (in minutes), during which each auxiliary fan was operating. The “Fan Summary” macro was then used to divide the total mining activity simulation time frame (in this scenario 250 hours) into 30-minute intervals and within each interval to determine and display which of the auxiliary fans were operating. Through the use of the “Fan Volume” worksheet the same macro assigned the air volumes (as defined in section 6.2.2) to the operating fans, then calculated and displayed the combined activity based air volumes required by the mining block versus traditional ventilation practice, in table and graphical format (see Figure 25). 97  Figure 20: VRM 1 production activity durations and occurrence times (in AC)  98  Figure 21: VRM 2 production activity durations and occurrence times (in AC)  99  Figure 22: Development 1 and Development 2 activity durations and occurrence times (in AC)  100  Figure 23: Fan activity start and end times in hours (absolute clock) for VRM 1 and VRM 2  101  Figure 24: Fan activity start and end times in hours (absolute clock) for Development 1 and Development 2  102  The first graph in Figure 25 shows the number of auxiliary fans that were active at any one time during the simulation time frame, in this case 250 hours (10.5 days). This graph shows that of the total number of 12 auxiliary fans that were assigned to the mining block, the maximum number of fans that were concurrently “On” at any time, was 10. Furthermore, this graph shows that during a large portion of the simulation time frame, only 6 auxiliary fans were concurrently operating. The second graph from Figure 25 shows the total air volume that is required by the mining block for activity based versus traditional ventilation. With activity based ventilation and diesel exhaust dilution considerations, the “Fan Summary” macro has determined that during this simulation time frame, on average, the total fresh air required by the mining block was: QActivity-30min. = 77.3 m3/s. Based upon traditional ventilation practice and diesel exhaust considerations, the same “Fan Summary” macro determined that total fresh air required by the mining block was: QTraditional = 114.7 m3/s. This shows that with activity based ventilation, the total fresh air required by the mining block is ΔQ30min. = 37.4 m3/s, or 33% less than the total fresh air required by the mining block with traditional ventilation practice. The third graph from Figure 25 shows the combined air volume that was “handled” by all auxiliary fans during the simulation time frame. As previously mentioned, under some operating circumstances such as during ore haulage operations that immediately follow a blasting fume clearance time period, to maintain adequate working conditions for men and machinery, the auxiliary fans responsible for providing fresh air to the draw-points may be required to be turned “On” even before the LHD enters the production area. For these particular fans, to account for additional fan operating times, the user would need to provide the “early start” times, which can be entered using the “Fan Volume” worksheet (see Figure 17). Furthermore, within the “Fan Summary” macro interface, the user would need to select the “Fan Start Early” text box and provide the row number of the “Fan Volume” worksheet where the early fan(s) start times were entered (see Figure 19). The “Fan Summary” macro would then re-divide the simulation time frame, re-assign the air volumes to the active fans and calculate the combined air volume “required” by the mining block. For example, with the early fan start times (as entered in Figure 17), the number of operating fans, the combined air volume “required” by the mining block and the combined air volume “handled” by the auxiliary fans are shown in Figure 26. For the same “Custom Report” data generated by AutoModTM model runs as shown in Figure 23 & Figure 24, the “Fan Summary” macro was again used to divide the mining process 103  simulation time frame (i.e. 250 hours) into much narrower time intervals (i.e. 5-minute intervals). Within each of these intervals the “Fan Summary” macro determined and displayed the operating fans. Subsequently, the “Fan Summery” macro assigned the predetermined air volumes (as defined in section 6.2.2) to the auxiliary fans and computed the combined air volume “required” by the mining block for activity based versus traditional ventilation in tabular and graphical formats (see Figure 27). The first graph from Figure 27 shows the number of fans that were operating within each of the 5-minute time intervals. When compared to Figure 26, this fan graph shows that by narrowing the time interval from 30 minutes to 5 minutes, the number of operating fans is decreasing. This would actually be expected as the “Fan Summary” macro would identify fewer concurrently operating fans, as the mining process simulation time frame is divided into much narrower time intervals, namely from 30-minute to 5-minute intervals. In other words, by dividing the mining process simulation time frame into very fine time intervals, the combined air volume calculated by the “Fan Summary” macro would converge towards “true” activity based air volume requirements. For example, the second graph from Figure 27, shows that when dividing the mining process simulation time frame into 5-minute intervals the average combined air volume “required” by the mining block reduces to 60.9 m3/s, which indicates air volume savings of ΔQ5-min. = 53.8 m3/s, a 47% reduction. However, this would probably represent an ideal operating scenario where the On/Off status of all auxiliary fans is purely triggered by the arrival/departure of the mining machinery to/from the control points located along the haulage drifts. A more realistic activity based air volume demand schedule is shown in Figure 28, where the “Fan Summary” macro takes into account “early fan start” times for all mining activities that immediately follow stope/development blasting fume clearance time periods. With the mining process simulation time frame divided into 5-minute intervals and “early fan start” operating conditions accounted for, the combined activity based air volume “required” by the multi-level mining bloc was determined at QActivity = 73.0 m3/s. Again, with traditional ventilation practice, the combined air volume required by the mining block was: QTraditional 114.7 m3/s. This indicates air volume savings of ΔQActivity = 41.7 3/s, a 36% intake flow reduction.  104  Figure 25: Air volume required at the mining block (with simulation time divided into 30-minute intervals)  105  Figure 26: Air volume required at the mining block (with simulation time divided into 30-minute intervals and early fan start)  106  Figure 27: Air volume required at the mining block (with simulation time divided into 5-minute intervals)  107  Figure 28: Air volume required at the mining block (with simulation time divided into 5-minute intervals and early fan starts)  108  6.1.5 Summary - Output Data Manipulation Mining process simulations performed on the “Mining Process” model showed that the model initiation functions, the process logic, the mining activity and ventilation variables had the ability to dynamically display in “real time” as well as in “accelerated” or “decelerated” animation environments the mining processes and fan activities based upon the mining operation schedules (e.g. drilling, explosive loading, blasting, ore/waste transportation) entered by the user. The ventilation output data generated through mining process simulation could be analyzed and processed into activity driven air volume schedules based upon the mining operations that were performed within the mining block. For the “Mining Process” model, the stope and development capacities, equipment characteristics and mining activity schedules were provided through the use of a spreadsheet utility, namely the “Data Input” utility. From the “Data Input” utility, mining activity schedules were then transferred to the model initiation functions and the process logic of the simulation model by a visual basic application. Furthermore, the traditional and activity based air volume requirements for the mining block were determined from ventilation output data generated by AutoModTM simulation. Within the “Mining Process” model, the ventilation parameters, such as the air volumes associated with the auxiliary fans were entered through the “Fan Volume” worksheet utility that was developed inside the “Results” data file. For a specified simulation time frame and based upon the scheduled mining operations entered through the “Data Input” utility, the ventilation variables defined within the process logic of the simulation model were able to determine when and for how long each of the auxiliary fans assigned to the mining block were “On”, information which was then transferred and printed in the “Custom Report” worksheet of the “Results” data file. The “Fan Summary” macro was then used to divide the total simulation time-frame into varying time intervals (from 1 to 60 minutes) that could be selected by the user, and for the selected time interval to determine and display the operating fans. Furthermore, the “Fan Summary” macro was then used to associate the predefined air volumes with the operating fans and calculate the combined activity based air volume needed by the mining block.  109  For model validation purposes, during a predetermined simulation time frame (e.g. production shift), the duration of the mining operations (drilling, explosive loading, blasting, mucking), their occurrence time and their completion sequence, which were dynamically displayed by AutoModTM simulator were compared with mining activity data collected within a multi-level mining block that employs the VRM method. The mining block’s activity based intake air volume requirement determined by means of process simulation, where the variable air volumes delivered to the production workings are managed by a yet to be installed ventilation control system, could also be validated through the use of a test stope as follows: •  Ultrasonic airflow sensors would be installed along the drilling level, the mucking level and around the draw-points of the test stope.  •  An ultrasonic airflow sensor would be installed inside each auxiliary duct through which fresh air is delivered to the drilling site and to the draw-points of the test stope.  •  An ultrasonic airflow sensor would also be installed within the immediate vicinity of the RAR through which the contaminated air coming from the drilling site and the draw-points is directed into the mine’s return air system.  •  Each airflow sensor would be equipped with a data logger. At the end of the production shifts, these data would be downloaded onto a mobile computer and processes into real-time air volume trends.  •  These air volume trends would then be compared with the air volume trends generated by discrete-event process simulations performed on the “Mining Process” model using AutoModTM.  110  7.  DETERMINATION OF SHORT-TERM AND LIFE-CYCLE AIRFLOW REQUIREMENTS FOR METAL MINES  The case studies presented in sections 5.3 and 5.4 have shown that determining the primary air volume requirements in large and deep mechanized metal mines is not straightforward due to the dynamic nature of the production processes that constantly redefine where air is needed. The case study presented in section 5.4 showed that based upon production planning data alone the mine’s intake air volume requirement could be greatly underestimated unless provisions in additional air volumes are considered in order to prevent uncontrolled recirculation at the pick-up location of the auxiliary fans and to maintain minimum airflow velocities throughout the non-productive levels. Further to this, the use of the mine’s diesel equipment activity data from its SIMS database has captured the ever changing and dynamic nature of production and development processes, however this historical data seemed to overestimate the mine’s intake air volume requirement due to insufficient information with respect to when, where and for how long the mining equipment was operating. The case study presented in section 5.3 also highlighted the lack of sufficient information with respect to where and for how long production and development equipment was operating. As a result, to determine the airflow requirements within the mine’s auxiliary ventilation system was problematic due to a large number of auxiliary fans (>200 fans) that were employed by the independent auxiliary systems delivering fresh air to the active and temporarily inactive mining blocks. The primary ventilation system is a major contributor to the capital and operating cost of most mines. It also has a major bearing on the health and safety of the underground workforce. Probably the most important single design parameter for the primary ventilation system is to accurately determine its overall intake air volume required by a mine [Brake & Nixon, 2008].  7.1  Determination of Short-Term Airflow Requirements  Using the “Custom Report” data generated by AutoModTM model runs as shown in Figure 23 and Figure 24, the “Fan Summary” macro determined the activity based air volume requirements based upon the capacity of the auxiliary fans that were assigned to the 111  independent auxiliary systems of the multi-level mining block. With the “early fan start” operating condition imposed on all auxiliary fans that delivered flow for mining operations that immediately followed stope & development blasting fume clearance times, the “Fan Summary” macro determined that the activity based air volume requirement at the multilevel mining block (presented in section 6.2.4) was: Qactivity-based = 73.0 m3/s versus the air volume required at the same mining block but based upon traditional ventilation, namely Qtraditional = 114.7 m3/s. Within each active mining block, this represents air volume savings of 41.7 m3/s, or 36%. These air volume savings generated within the mining blocks auxiliary ventilation system are then further translated in reductions of the mine’s total intake airflow.  7.1.1 Short-Term Airflow Requirements – Traditional Ventilation Practice With traditional ventilation practice the required air volumes for the development and production workings of the mining block were determined according to the diesel exhaust dilution criteria of 0.063 m3/s/kW (100cfm/bhp) that was applied to the maximum number of diesel units that could operate within these working areas at any time (see section 6.2.2). With traditional ventilations practice and based upon the diesel exhaust dilution criteria, the auxiliary fans that are assigned to the mining block would need to deliver the following air volumes to the development and production workings: •  No. 1 fan located inside the auxiliary steel duct installed along the VRM #1 drilling level: Q(Scav1-drill) = 12.0 m3/s  •  No. 2 and No. 3 fans located inside the auxiliary steel ducts installed along the VRM #1 mucking level: − Q(Fan1-scoop) = 21.3 m3/s − Q(Scav-fan1-scoop) = 18.0 m3/s  •  No. 4 fan located inside the auxiliary steel duct installed along VRM #2 drilling level fan: Q(Scav2-drill) = 12.0 m3/s  •  No. 5 and No. 6 fans located inside the auxiliary steel ducts installed along the VRM #2 mucking level: − Q(Fan2-scoop) = 21.3 m3/s − Q(Scav-fan2-scoop) = 18.0 m3/s  •  No. 7 fans located inside the auxiliary steel duct installed along Ore-pass #1 haulage drift: Q(Orepass 1) = 21.3 m3/s 112  •  No. 8 fan located inside the auxiliary steel duct installed along Ore-pass #2 haulage drift: Q(Orepass 2) = 21.3 m3/s  •  No. 9 fan located inside the auxiliary steel duct installed along Development #1: Q(Dev1-fan-scoop) = 14.8 m3/s  •  No. 10 fan located inside the auxiliary steel duct installed along Development #2: Q(Dev2-fan-scoop) = 14.8 m3/s  •  No. 11 fan located inside the auxiliary steel duct installed along Waste-pass #1 haulage drift: Q(Dev1-fan) = 14.8 m3/s  •  No. 12 fan located inside the auxiliary steel duct installed along Waste-pass #2 haulage drift: Q(Dev2-fan) = 18.5 m3/s  Based upon these air volumes delivered by the auxiliary fans, for traditional ventilation practice, the total air volume required by this multi-level mining block is Qblock-traditional = 115 m3/s. The airflow distribution along the access ramp, VRM stopes, developments and along the haulage drifts that connect the production and development workings to the ore/waste passes of this mining block is shown in Figure 29. One method to calculate an underground operation’s total intake air volume can be based upon the maximum number of active stopes that are needed to achieve short or long-term production targets. Within the next sections, for practical purposes, the total intake air volume was determined for an existing multi-level metal mine, namely “Mine A”, which employs the vertical retreat mining method to extract the ore reserves from a steeply deeping orebody. Besides the fresh air volumes required by the mining blocks, the mine’s total intake airflow includes the air volumes required by all secondary services such as backfilling, ore crushing and loading operations, equipment maintenance (garage) as well as minimum air volumes needed within the inactive areas of the mine. For year 2008 and based upon the number of active stopes (or mining blocks) that are needed to achieve production targets, Mine A’s total intake air volume can be determined as follows: •  Daily production requirement: 4,000 tons/day  •  Broken ore transported from one VRM stope (ST8 Wagner LHD): 500 tons/day  •  Broken ore transported from one mining block: 500 tons/stope x 2 stopes/mining block = 1,000 tons/day  113  •  Required active mining blocks: 4,000 tons/day/ 1,000 tons/day = 4 mining blocks (or 8 active stopes)  •  Air volume required by the mining blocks: Qproduction-traditionl = 4 blocks x 115 m3/s = 460 m3/s  •  Air volume needed to maintain minimum flow velocity within the inactive areas of the mine (within the upper levels): Qupper-levels = 100 m3/s  •  Air volume required by the garage: Qgarage = 15 m3/s  •  Air volume required for backfill operations: Qbackfill = 30 m3/s  •  Air volume required at the crushing station: Qcrusher = 20 m3/s  •  Air volume required at the loading pockets: Qpockets = 15 m3/s  •  Air volume required to maintain minimum flow velocity along the access ramp: Qramp-min = 20 m3/s  At “Mine A”, for year 2008 the total air volume required by secondary services and to maintain minimum velocities within the primary ventilation system is: Qservices = 100 m3/s + 15 m3/s + 30 m3/s + 20 m3/s + 15 m3/s + 20 m3/s = 200 m3/s. As a result, for traditional ventilation practice, the total intake air volume required at “Mine A” is: QIntake-traditional = Qproduction-traditional + Qservices = 460 m3/s + 200 m3/s = 660 m3/s. The mine’s total air volume and the airflow requirements for production and secondary mining operations are also presented in a schematic format (see Figure 30).  114  Figure 29: Airflow distribution within the mining block – traditional ventilation practice  115  Figure 30: Airflow requirements for production and secondary activities – Year 2008 (Traditional Ventilation Practice)  116  7.1.2 Short-Term Airflow Requirements – Activity Based Ventilation For activity based ventilation practice, the fresh air required by one mining block with two VRM stopes, two developments and their associated ore and waste passes was determined through mining process simulation using AutoModTM, as shown in section 6. Based upon the “Custom Report” data generated by AutoModTM model runs, the “Fan Summary” macro had the ability to determine and display which of the auxiliary fans assigned to the mining block were operating at any one time within the simulation time frame. As described in section 6.2.4, the “Fan Summary” macro showed that from the total numbers of 12 auxiliary fans assigned to the multi-level mining block, the maximum number of auxiliary fans that were ON at any one time was 9. This graph also showed that during most of the simulation time frame only 6 of the auxiliary fans were concurrently ON (see Figure 28). With the simulation time frame divided into 5-minute intervals, the “Fan Summary” macro determined that the “true” activity-based air volume required at the mining block is: Qblock-activity = 73.0 m3/s. Again, for year 2008 the total air volume required by secondary services and to maintain minimum velocities within the inactive areas of the mine is: Qservices = 100 m3/s + 15 m3/s + 30 m3/s + 20 m3/s + 15 m3/s + 20 m3/s = 200 m3/s. The required air volume by the mining blocks is: Qproduction-activity = 4 mining blocks x 73 m3/s = 292 m3/s. As a result, for activity based ventilation, the total fresh air required by “Mine A” is: QIntake-activity = Qproduction-activity + Qservices = 292 m3/s + 200 m3/s = 492 m3/s. This represents savings of 168 m3/s (25%), when compared to the total intake air volume determined for traditional ventilation practice. For activity based ventilation practice, Figure 31 shows the mine’s total intake air volume as well as the airflow requirements for production and secondary mining operations in a schematic format.  117  Figure 31: Airflow requirements for production and secondary activities – Year 2008 (Activity Based)  118  7.2  Determination of Life-Cycle Airflow Requirements  The two case studies that were presented in sections 5.3 and 5.4 have shown that determining long-term air volume requirements for large and deep metal mines by means of conventional ventilation planning techniques can be very difficult and extremely labour intensive. The first case study showed that determining the air volume requirements for the production and development workings was problematic due to a large number of auxiliary fans (>200 fans) and lack of historic data as to when and for how long these fans were operating. The second case study showed again that determining long-term airflow was not straightforward due to the dynamic nature of production and development activities constantly redefining where and how much airflow is required in order to provide safe environmental conditions for the underground workforce. In the second case study two airflow evaluation methods were used to determine the potential ventilation redundancy at the mine. The first method was based upon the mine’s future production planning data to determine the mines long term airflow requirement. The second method was based upon retrospective equipment activity data collected from the mine’s SIMS database. This case study showed that both methods had limitations in determining the mine’s longterm air volume requirement schedule: •  The “future production planning” analysis (2nd stage review) underestimated the mine’s total intake air volume because no additional air volume was allocated in order to prevent uncontrolled airflow recirculation within the ventilation system and provide minimum air volumes to the inactive levels of the mine. Even with this provisions in place, the “future production planning” analysis continued to underestimate the mine’s intake airflow due to insufficient information with respect to where production and development activities were taking place.  •  The “retrospective equipment activity” analysis (3rd stage review) had the ability to capture the dynamic nature of the mine. However, this data was still insufficient with respect to where mining activities were taking place and for how long. As a result, tis airflow evaluation method seemed to overestimate the mine’s total intake airflow requirement. This was mainly due to the fact that the mining equipment operating in the production workings did not have a “time stamp”. Consequently, fresh air was 119  assigned multiple times to the same mining equipment entering a production level but at different times.  7.2.1 Life-Cycle Airflow Requirements – Traditional Ventilation Practice For traditional ventilation practice, the total intake air volume required by “Mine A” for its operating life can be determined by summing up the total planned air volume needed by the mining blocks for production and development activities (Qproduction-traditional) and the total air volume required by secondary services (Qservices). Again, the air volume required by secondary services also includes provisions in air volume that are needed to maintain minimum flow velocities within the inactive areas of the mine. The total planned air volume required by the development and production workings is dictated by the number of mining blocks that are needed to be maintained active in order to fulfill production requirements. For the remaining life of “Mine A”, long-term planning data show the following forecasted production requirements: •  Between years 2008 – 2010: P2008-2010 = 4,000 tons/day  •  Between years 2011 – 2013: P2011-2013 = 5,000 tons/day  •  Between years 2014 – 2017: P2014-2017 = 6,000 tons/day  •  Between years 2018 – 2020: P2018-2020 = 5,000 tons/day  At “Mine A”, the amount of broken ore produced from a mining block is 1,000 tons/day. As a result, for the above mentioned time periods the number of mining blocks that are needed to be active are: •  Between years 2008 – 2010: N2008-2010 = 4 mining blocks (8 VRM stopes)  •  Between years 2011 – 2013: N2011-2013 = 5 mining blocks (10 VRM stopes)  •  Between years 2014 – 2017: N2014-2017 = 6 mining blocks (12 VRM stopes)  •  Between years 2018 – 2020: N2018-2020 = 5 mining blocks (10 VRM stopes)  Figure 32 shows the long-term forecasted production rates and the required active mining blocks in a graphical format.  120  Figure 32: Forecasted production requirements for the remaining operating life of “Mine A”  Forecasted Production Rates & Active Mining Blocks at Mine C 7000 6,000 tons/day  Production Rate (tons/day)  6000 5,000 tons/day  5,000 tons/day  5000 4,000 tons/day 4000  3000 4 Mining Blocks (8 VRM Stopes)  5 Mining Blocks (10 VRM Stopes)  6 Mining Blocks (12 VRM Stopes)  5 Mining Blocks (10 VRM Stopes)  2000  1000 Period: 2008 - 2010  0 2008  2009  2010  Period: 2011 - 2013  2011  2012  2013  Period: 2014 - 2017  2014  Year  2015  2016  Period: 2018 - 2020  2017  2018  2019  2020  121  Assuming that during its operating life, the total air volume required by secondary services remains the same, for traditional ventilation practice, the life-cycle air volume requirements at “Mine A” are: •  Between years 2008 – 2010: QIntake(2008-2010) = Qproduction-traditional + Qservices = 4 mining blocks x 115 m3/s/block + 200 m3/s = 460 m3/s+200 m3/s = 660 m3/s  •  Between years 2011 – 2013: QIntake(2011-213) = Qproduction-traditional + Qservices = 5 mining blocks x 115m3/s/block = 575 m3/s+200 m3/s = 775 m3/s  •  Between years 2014 – 2017: QIntake(2014-2017) = Qproduction-traditional + Qservices = 6 mining blocks x 115m3/s /block = 200 m3/s = 690 m3/s+200 m3/s = 890 m3/s  •  Between years 2018 – 2020: QIntake(2018-2020) = Qproduction-traditional + Qservices = 5 mining blocks x 115m3/s/block = 575 m3/s+200 m3/s = 775m3/s  7.2.2 Life-Cycle Airflow Requirements – Activity Based Ventilation Based upon the “Custom Report” data generated through discrete-event process simulation performed on the “Mining Process” model using AutoModTM, the “Fan Summary” macro determined that the “true” activity based air volume required by the multi-level mining block is: Qblock-activity = 73.0 m3/s. Again, assuming that during its operating life, the total air volume required by secondary services remains the same, for activity based ventilation practice, the life-cycle air volume requirements at “Mine A” are: •  Between years 2008 – 2010: QIntake(2008-2010) = Qproduction-activity + Qservices = 4 mining blocks x 73.0 m3/s /mining block + 200 m3/s = 492 m3/s  •  Between years 2011 – 2013: QIntake(2011-213) = Qproduction-activity + Qservices = 5 mining blocks x 73.0 m3/s /mining block + 200 m3/s = 565 m3/s  •  Between years 2014 – 2017: QIntake(2014-2017) = Qproduction-activity + Qservices = 6 mining blocks x 73.0 m3/s /mining block + 200 m3/s = 638 m3/s  •  Between years 2018 – 2020: QIntake(2008-2010) = Qproduction-activity + Qservices = 5 mining blocks x 73.0 m3/s /mining block + 200 m3/s = 565 m3/s  122  Figure 33 shows the life-cycle airflow demand schedules at “Mine A” for traditional versus activity based ventilation practices in a graphical format. This graph shows that with traditional ventilation practice the current air volume requirements at “Mine A” will remain the same until year 2010. In 2011, due to increased production requirements, namely from 4,000 to 5,000 tons/day, the mine’s total intake air volume will increase from 660 m3/s to 775 m3/s, which represents an increase of 115 m3/s (or 17%). Between years 2014 – 2017, production requirements will further increase from 5,000 tons/day to 6,000 tons/day. Consequently, the mine’s total intake air volume will also increase to 890 m3/s. Preliminary ventilation simulations performed on a ventilation model representing this future configuration of “Mine A”, showed that this intake air volume of 890 m3/s is well beyond the current capacity of the mine’s primary ventilation system. To deliver this fresh air to the development and production workings and for secondary services, in addition to the existing ventilation infrastructure a new intake FAR and additional primary fans are needed. The economic impact of this future ventilation infrastructure will be discussed in section 9. Figure 33 also shows that with activity based ventilation practice, during its operating life, the mine’s intake air volume savings are: •  Between years 2008 – 2010: ΔQIntake(2008-2010)  = 168 m3/s, which represents  approximately 25% less intake air volume as for traditional ventilation practice •  Between years 2011 – 2013 and 2018 - 2020: ΔQIntake(2011-213) = 210 m3/s, which represents approximately 27% less intake air volume, and  •  Between years 2014 – 2017: ΔQIntake(2014-2017) = 252 m3/s, which represents 28% less intake airflow  At “Mine A”, airflow calculations have shown that during its operating life, with an activity based auxiliary ventilation system that delivers appropriate air volumes when and where mining activities occur, savings in intake air volumes can range from 25% to 28%. Consequently, due to the cubic relationship between the fan power and the mine’s total intake air volume (see equation 3 in section 1.1.1), the operating cost of the mine’s primary and auxiliary ventilation systems can be significantly reduced. The economic and environmental benefits of this new ventilation design approach will be presented and discussed in section 9.  123  Figure 33: Life-Cycle Airflow demand Schedule for “Mine A” – Traditional vs. Activity Based Ventilation Practice  Life-Cycle Airflow Demand Schedule for Mine C Traditional vs. Activity Based Ventilation Practice 1000 890 m3/s 900  Air Volume (m3/s)  700  775 m3/s  775 m3/s  800 660 m3/s  638 m3/s 565 m3/s  565 m3/s  600 492 m3/s 500 400 300  Δ Q = 168m3/s (25%)  Δ Q = 210m3/s (27%)  Δ Q = 252m3/s (28%)  Δ Q = 210m3/s (27%)  200 100  Period: 2008 - 2010  Period: 2011 - 2013  Period: 2014 - 2017  Period: 2018 - 2020  0  2008  2009  2010  2011  2012  2013  2014  2015  2016  2017  Years Traditional Ventilation  Activity Based Ventilation  2018  2019  2020  124  8.  VENTILATION SYSTEM DESIGN BASED UPON THE LIFECYCLE AIRFLOW DEMAND SCHEDULE  In general, almost every underground mine is unique in its orebody geometry, geological characteristics and potential environmental pollutants. As a result, the airflow distribution throughout the mines’ primary and auxiliary ventilation systems can be quite variable. However, underground metal mines have certain essential ventilation characteristics that are common, such as: •  Usually fresh air enters the mines through one or more shafts or vertical airways and near vertical ventilation raises connected to surface  •  The fresh air flows down the intake airways to the working areas where majority of pollutants are added to the airflow  •  The contaminated air is directed along the return airways to the return system consisting on one or more upcast shafts and/or vertical & near-vertical airways  In underground mines the primary ventilation system includes all vertical, near-vertical and horizontal airways through which fresh air is delivered to the production areas as well as the airways through which contaminated air is directed to surface. From the level or transfer drifts, the fresh air is picked up and delivered to individual production drawpoints or development faces by the auxiliary ventilation system. The most common auxiliary ventilation system used in Canadian metal mines is the fan-duct auxiliary system in its “forcing” or “exhausting” setups (see section 2.1 for details). Before analyzing the potential benefits of this new design approach, the primary and auxiliary ventilation systems of “Mine A” need to be first solved and balanced. The first step in designing a mine’s ventilation system is to establish the air volumes required in all places of work, haulage, travel or other activity areas such as crushing stations, ore and waste loading pockets. In addition to areas of rock fragmentation, in an underground operation there are other subsurface facilities that require the environment to be controlled. These include workshops, stationary equipment such as pumps or electrical sub-stations, fuel stations and storage areas. Furthermore, additional air volumes need to be considered in order to maintain minimum airflow velocities within the inactive areas of the mine. Within this phase of the ventilation system design, airflow velocity limits are also determined along the main airways. The recommended maximum airflow velocities within the working 125  faces are between 1 m/s to 4 m/s. Discomfort can be experienced by personnel at velocities in excess of 4 m/s because of impact by large dust particles, particularly in cool conditions. On the other hand, excessive airflow velocities within the mine can also result in unacceptable ventilation operating costs [McPherson, 1993]. The dynamic nature of an underground operation requires that the infrastructure of the ventilation system be designed such that it can accommodate major changes during its operating life. These major changes can occur when: •  More ore reserves are to be extracted  •  Production and consequently ventilation requirements increase to an extent that the existing primary fans and the ventilation infrastructure are incapable of providing the required face airflows at an acceptable operating cost  To accommodate major changes during the operating life of a mine, a “time-phase” exercise can be carried out on ventilation models that represent the configuration of the mine for each time period that incorporates the above mentioned changes. This time- phase exercise can be conducted to cover the life of the mine, or as far into the future as can reasonably be predicted. For example, with traditional ventilation practice, preliminary simulations performed on Mine A’s ventilation model show that due to increased production requirements between years 2014 – 2017, the current primary fans will be unable to provide the required air volumes through its current ventilation infrastructure and it appears inevitable that a new fresh air system will be required. It would probably be prudent to first investigate the time-phase when this new intake fresh air raise has become necessary and then examine earlier time-phases as well. This would determine whether it is more advantageous to construct the fresh air raise at a prior time rather than wait until it becomes absolutely necessary. Furthermore, it may be extremely cost effective to investigate alternative solutions, such as an activity based ventilation system that would reduce the mine’s total intake air volume by eliminating unnecessary ventilation within its production areas. The life-cycle airflow demand schedule presented in section 7.2 showed that between years 2014 – 2017, the total intake air volume for an activity based ventilation practice at “Mine A” is 638 m3/s, which is less than what the mine currently supplies, namely 660 m3/s. Consequently, for its remaining life, Mine A’s primary ventilation infrastructure would not require any major upgrades.  126  As previously mentioned, to determine the potential benefits of an activity based ventilation system designed according to the life-cycle airflow demand schedule that was determined through AutoModTM simulations, the primary and auxiliary ventilation systems of “Mine A” need to be solved and balanced for different time periods where the mine’s air volume requirements significantly change. The long-term production planning data showed that during its remaining life the production requirements at “Mine A” vary between 4,000 to 6,000 tons/day. Consequently, the number of mining blocks that would need to be equipped with auxiliary ventilation systems and be maintained active will also vary. Based upon the life-cycle air volume demand schedule presented in Figure 33, Mine A’s primary and auxiliary systems for both traditional and activity based air volume requirements will be solved and balanced for the following time periods: •  Years (2008 – 2010)  •  Years (2011 – 2013)  •  Years (2014 – 2017), and  •  Years (2018 – 2020)  For the first time-period, namely years 2008-2010, for both traditional and activity-based air volume requirements, the fan operating duties, the airflow distribution within the primary and anxiety systems, the fan power and operating costs determined through ventilation simulation will be presented in the main body of the thesis (section 8.1.1 and section 8.2.1). For the subsequent time-periods, namely years 2011 – 2013, 2014 – 2017 and 2018 – 2020, for both traditional and activity based air volume requirements, a detailed description of the ventilation modelling work and ventilation simulation results are presented in “Appendix B” (section B.1.1 through to section B.1.3 and section B.2.1 through to section B.2.3).  8.1  Solving and Balancing the Primary and Auxiliary Ventilation Systems of “Mine A” for Years 2008–2010 and for Traditional Air Volume Requirements  8.1.1 Ventilation Modelling for Years 2008–2010 (Traditional Ventilation) Presently, “Mine A” produces 4,000 tons/day from a near continuous operation comprising a schedule of two back-to-back 10-hour shifts per day, with a 4-hour blast clearance period during the night. The mining method employed at this mine is vertical retreat mining. The ore is extracted from 4 mining blocks located within the bottom levels of the mine, namely 127  between the 7200 and 7900 levels (See Figure 34). The mining blocks have similar geometrical characteristics. Each mining block includes two developments and two VRM stopes as follows: •  Mining block No. 1 with stopes on the 7200 and 7400 levels  •  Mining block No. 2 with stopes on the 7530 and 7660 levels  •  Mining block No. 3 with stopes on the 7720 and 7780 levels, and  •  Mining block No. 4 with stopes on the 7840 and 7900 levels  Each production stope has a drilling and a mucking level. The mucking level has 8 drawpoints from where the ore is transported to the ore-pass. The drilling and mucking levels have their own independent auxiliary ventilation system. The upper orebody of “Mine A” has three main intake airways; No. 9 production shaft from surface to the 7000L (2,100m), the main fresh air raise (Main FAR) and the ramp system, through which fresh air is delivered to the 7200L (2,200m). In addition the upper orebody has an internal shafts (No. 8 Shaft) through which fresh air is delivered from the 5000L (1,500m) to the 7200L (2,200m). The deep orebody has a single main intake fresh air raise (Main FAR) and an extension that will deliver fresh air to the future bottom level of the mine, namely to the 8,200L (2,500m). The mine’s exhaust system includes the deep orebody’s main return air raise (Main RAR) and the No. 11 exhaust shaft. Primary System – Flow Distribution, Fan Duties and Power Requirements To deliver the required air volume down the intake airways and along the main level drifts the mine employs a push-pull primary ventilation practice. Presently, in its primary system the mine uses 13 primary and booster fans, the majority of which (ten fans in parallel arrangement of two) are located in the intake system. The remaining 3 fans, also in parallel arrangement are located on top of the No. 11 exhaust shaft. For the mining blocks, fresh air is drawn off of the main intake system and is delivered to the stope drawpoints and to the face of the developments through independent auxiliary ducting systems under the assistance of the auxiliary fans. From the production and development workings the contaminated air is exhausted to surface via the main RAR and No. 11 exhaust shaft. Within its 4 active mining blocks the mine employs 48 auxiliary fans with motor 128  powers ranging between 25 to 75 kW. To provide the air volumes required by the mine’s secondary services and fixed infrastructures (i.e. garage, crusher, loading pockets) as well to provide minimum airflow velocities throughout the inactive upper levels the mine also uses 62 auxiliary fans with motor powers ranging between 25 to 75 kW. For the 2008 – 2010 time period, the ventilation model of Mine A’s primary system was developed based upon ventilation survey data and mine plans provided by its engineering department, using VnetPCTM software. For this time period, the mine’s total intake airflow (see section 7.2.1), which includes the air volumes needed for production and secondary services was determined at: QIntake(2008-2010) = Qproduction-traditional + Qservices = 460 m3/s = 200 m3/s = 660 m3/s. Based upon this intake air volume, the primary ventilation system of “Mine A” was solved and balanced using ventilation simulation techniques. The airflow distribution within the mine’s primary ventilation system is shown in Figure 34. Ventilation simulations also show that of its total intake air volume of 660 m3/s, 176 m3/s is delivered down the No. 9 shaft, 286 m3/s is delivered down the Main FAR, 183 m3 down the No. 8 internal shaft and 15 m3/s down the No. 3 shaft. Most of the contaminated air resulting from production and secondary mining activities, namely 644 m3/s is directed to surface through the No. 11 exhaust shaft. Approximately 16 m3/s of contaminated air resulting from secondary activities within the 3800 level is directed to surface through the No. 5 exhaust shaft.  129  Figure 34: Airflow distribution and fan operating duties in the primary system for years 2008 – 2010 - Traditional ventilation  130  Within the primary system the operating duties of the main fans were determined by means of ventilation simulation, as follows: •  1420 Level Intake Fans (2 fans in parallel): Q = 177 m3/s, Ptotal = 1,538 Pa  •  2300 Level Intake Fans (2 fans in parallel): Q = 286 m3/s, Ptotal = 2,136 Pa  •  2600 Level Intake Fans (2 fans in parallel): Q = 183 m3/s, Ptotal = 2,041 Pa  •  No. 11 Shaft Exhaust Fans (3 fans in parallel): Q = 644 m3/s, Ptotal = 3,440 Pa  The operating cost of the ventilation system can be defined by equation (7) as follows: OC = Fp/ηf x OD x OH x EC ($/year) Where:  (7)  Fp = total fan power of the ventilation system (kW) ηf = average efficiency of the fans (%) OD = number of operating days per year (days/year) OH = operating hours per day (hours/day) EC = average electricity cost per kWh ($/kWh)  Ventilation simulations also show that the total fan power within the primary ventilation system is: PPrimary-Traditional = 4,458.8 kW. Furthermore, for an average 70% fan efficiency and for $0.055/kWh electricity charge, the primary ventilation system operating cost is: OCPrimary (2008-2011)  = 4,458.8 kW/0.7 x 365 days/year x 24 hours/day x $0.055/kWh = $3,068,920/year.  The operating duties and the power of all primary fans (main and boosters) that operate within the primary ventilation system are shown in Table 12.  131  Table 12: Fan operating duties and power requirements for primary system (years 2008 – 2010) – Traditional ventilation Junction From 191 383 39 13 380 420 500 684 487 36 701 7721 7784 7845 682 684 7724 7781 7783 7849 53 232 9 6 TOTAL  Junction To 190 384 40 29 381 421 501 702 485 35 702 7720 7783 7844 683 34 7725 7782 7790 7850 54 58 10 5  Pressure (kPa) 0.070 0.244 2.093 1.638 1.094 2.910 0.224 0.169 0.108 0.867 0.425 0.249 0.163 0.165 0.062 0.072 0.065 0.085 0.017 0.040 1.538 2.136 2.041 3.440  Flow (m3/s) 6.3 10.7 286.0 183.0 19.1 3.0 8.5 15.6 19.0 5.8 36.6 12.4 55.6 58.2 4.8 14.2 21.7 24.2 14.7 40.1 177.0 286.0 183.0 644.0  Power (kW) 0.44 2.61 598.68 299.75 20.92 8.77 1.90 2.64 2.05 5.04 15.55 3.09 9.06 9.60 0.30 1.02 1.41 2.05 0.25 1.60 272.24 610.90 373.50 2,215.42 4458.80  Operating Cost Fan Curve Fan Fan ($/year) Status Arrangement Location 301 On 1 in Series 1900L 5 SHAFT 1,797 On 1 in Series 3800L 5 SHAFT 412,059 On 1 in Series 5400L Fans 206,316 On 2 in Parallel 5000 Fans (2para) 14,397 On 1 in Series 3800L Booster Fan 6,035 On 1 in Series 4200L Booster Fan 1,307 On 1 in Series 5000L Booster Fan 1,819 On 1 in Series 6800-7000 R booster 1,411 On 1 in Series 4200 access booster 3,468 On 1 in Series FAR booster 10,705 On 1 in Series 7000L booster 2,127 On 1 in Series 7720L booster 6,236 On 1 in Series To 7780L booster 6,606 On 1 in Series To 7840 booster 203 On 1 in Series 6800 Ra drift fan 702 On 1 in Series 6800 - 6600 ramp 969 On 1 in Series FAR to ramp 1,414 On 1 in Series FAR to 7780L 172 On 1 in Series 7720L -7780L fan 1,104 On 1 in Series 7840-7900 FAR fan 187,378 On 2 in Parallel 1420L Fans (2p) 420,474 On 2 in Parallel 2300L Fans (2p) 257,076 On 2 in Parallel 2600L Fans (2p) 1,524,844 On 3 in Parallel 11 Shaft fans (3p) 3,068,920  132  Auxiliary System – Flow Distribution, Fan Duties and Power Requirements With traditional ventilation practice the air volume requirements within the mining blocks were determined according to diesel exhaust dilution considerations and the maximum number of diesel units that can operate in the production, development and ore/waste dumping areas at any time. At this mine, the mining equipment used in the production blocks and their associated air volumes are similar to the diesel fleet described in section 6.2.2. Based on this, the air volume requirements within one mining block are: •  For the drilling levels of both VRM stopes: Qdrilling = 12 m3/s  •  For the mucking levels of both VRM stopes: Qmucking = 21.3 m3/s  •  For both development headings: Qdevelopment = 14.8 m3/s  •  Along both ore-pass haulage drifts: Qore-pass = 21.3 m3/s  •  Along #1 waste-pass haulage drift: Qwaste-pass1 = 14.8 m3/s  •  Along #2 waste-pass haulage drift: Qwaste-pass2 = 18.5 m3/s  The air volumes in the mining blocks were also assessed for minimum airflow velocity requirements. In this respect, along the ore/waste haulage levels with no equipment present, the minimum air volumes are: •  Along the stope haulage levels: QHmin = 7 m3/s  •  Along the development haulage levels: QDmin = 7 m3/s  •  Along any section of the access ramp: QRmin = 14 m3/s  During scheduled production down-times (e.g. lunch breaks), to maintain these minimum air volumes throughout the mining block and to avoid airflow recirculation at the flow pick-up location of the auxiliary fans, the minimum air volume required by the mining block is: QBlock-min = 33.3 m3/s At this mine, within each mining block, both production drilling levels consist of an overcut that accommodates eight possible drilling locations. The fresh air to these drilling locations is delivered along a 38 inches diameter auxiliary duct under the assistance of one auxiliary. The contaminated air from the drill is directed along the haulage drift into the RAR. The working area of the production stopes includes 8 draw-points (2 x 4 layout), with 4 draw-points located on one side and 4 draw-points located on the other side (scavenger side) of the stope’s centre line. Both mucking levels have their own independent auxiliary system 133  consisting of steel ducts and two auxiliary fans. The fresh air to the draw-points area is delivered along a 48” auxiliary duct. From here, to each individual draw-point the fresh air is delivered along a 38” auxiliary duct. Depending upon the drawpoint from where ore is pulled, one or both auxiliary fans located inside the steel ducts could be operational. From the draw-points, the contaminated air flows back along the haulage drifts into the RAR. For each development heading, the required fresh is picked-up from the main access ramp and delivered along a 38” steel duct with the assistance of one auxiliary fan. From the heading, the contaminated air flows back along the development into the RAR. The ore and rock passes have their own independent auxiliary ventilation system as well. The fresh air to the ore and waste passes is delivered along 48” and 38” steel ducts, respectively, with the assistance of two auxiliary fans. Within the active mining blocks, the ore and waste passes does not have a direct connection to the RAR system. Consequently, the return air from the ore and waste passes is directed back along the haulage drifts into the main access ramp where it mixes with the fresh air delivered down the ramp. The ventilation model of the mining block’s auxiliary system was developed according to mine layouts and information provided by the mine using VnetPCTM program. Based upon the above mentioned flow requirements the auxiliary system was solved and balanced through ventilation simulation. For traditional ventilation practice, the airflow distribution throughout the auxiliary ducts and the mining block is shown in Figure 35.  134  Figure 35: Airflow distribution and fan duties within the mining block’s auxiliary ventilation system - Traditional ventilation  135  With traditional practice, ventilation simulations showed that to deliver the required air volumes to the production workings and along the haulage drifts to the ore and waste passes, the total air volume required by the mining block is: QTotal-block = 114.7 m3/s Simulations also show the auxiliary fans operating duties in terms of air volume (Q) delivered versus total pressure (Ptotal) that is needed to overcome the frictional pressure losses along the auxiliary ducts are: •  No. 1 auxiliary fan on VRM #1 drilling level: Q = 12 m3/s, Ptotal = 1,664 Pa  •  No. 2 auxiliary fan on VRM #1 mucking level: Q = 21.3 m3/s, Ptotal = 1,111 Pa  •  No. 3 fan on VRM #1 mucking level (scavenger side): Q = 18.0m3/s, Ptotal = 1,363 Pa  •  No. 4 auxiliary fan on VRM #2 drilling level: Q = 12 m3/s, Ptotal = 1,627 Pa  •  No. 5 auxiliary fan on VRM #2 mucking level: Q = 21.3 m3/s, Ptotal = 1,116 Pa  •  No. 6 fan on VRM #2 mucking level (scavenger side): Q = 18.0m3/s, Ptotal = 1,302 Pa  •  No. 7 auxiliary fan in Development #1: Q = 14.8 m3/s, Ptotal = 2,125 Pa  •  No. 8 auxiliary fan in Development #2: Q = 14.8 m3/s, Ptotal = 2,127 Pa  •  No. 9 auxiliary fan in Ore-pass #1 haulage drift: Q = 21.3 m3/s, Ptotal = 1,161 Pa  •  No. 10 auxiliary fan in Ore-pass #2 haulage drift: Q = 21.3 m3/s, Ptotal = 1,163 Pa  •  No. 11 aux. fan in Waste-pass #1 haulage drift: Q = 14.8 m3/s, Ptotal = 1,238 Pa  •  No. 12 aux. fan in Waste-pass #2 haulage drift: Q = 18.5 m3/s, Ptotal = 1,801 Pa  With traditional ventilation practice, the operating duties and the operating power for each auxiliary fan assigned to the mining block is shown in Table 13. This table shows that the combined auxiliary fan power within the mining block is PBlock = 299 kW.  136  Table 13: Auxiliary fans operating duty and power requirements within the mining block - Traditional ventilation Junction Junction Pressure Flow Power Operating Cost Fan Curve Fan From To (kPa) (m3/s) (kW) ($/year) Status Arrangement 93 92 91 90 73 60 43 40 32 16 13 6 TOTAL  Location  83 69 55 28 74 61 44 41 33 17 14  1.801 1.238 1.163 1.161 2.127 2.125 1.302 1.116 1.627 1.363 1.111  18.5 14.8 21.3 21.3 14.8 14.8 18.0 21.3 12.0 18.0 21.3  33.32 18.32 24.77 24.73 31.48 31.45 23.44 23.77 19.52 24.53 23.66  22,933 12,611 17,050 17,021 21,667 21,647 16,131 16,361 13,438 16,886 16,288  On On On On On On On On On On On  1S 1S 1S 1S 1S 1S 1S 1S 1S 1S 1S  Waste pass 2 aux. fan (dev2_fan-drill) Waste pass 1 aux. fan (dev1_fan_drill) Orepass 2 auxiliary fan (orepass 2) Orepass 1 auxiliary fan (orepass 1) Development 2 fan (dev2_fan_scoop) Development 1 fan (dev1_fan_scoop) VRM2 scavenger fan (scav_fan2_scoop) VRM2 mucking level fan (fan2_scoop) VRM2 drilling level fan (scav2_drill) VRM1 scavenger fan (scav_fan1_scoop) VRM1 mucking level fan (fan1_scoop)  7  1.664  12.0  19.97 299.0  13,744 205,777  On  1S  VRM1 drilling level fan (scav1_drill)  137  The long-term air volume schedule for years 2008 - 2010 determined in section 7.2.1 shows that within this time frame ore reserves are extracted from 4 mining blocks. Consequently, the total fan power in the production stopes is: PProduction = 299 kW/block x 4 mining blocks = 1,196 kW. Section 7.1.1 showed that fresh air is also required for secondary services such as backfilling, ore crushing, fixed infrastructures and to provide minimum flow to the inactive areas of the mine. To deliver these air volumes, the mine uses 62 additional auxiliary fans. The auxiliary ventilation setup for the secondary services is similar to the auxiliary ventilation setup within the mining blocks. The total fan power for the secondary services is: PServices = 24.9 kW x 62 fans = 1,544 kW. The combined fan power of the mine’s auxiliary system is: PAuxiliary = PProduction + PServices = 1,196 kW + 1,544 kW = 2,740 kW. For an average 70% fan efficiency and for $0.055/kWh electricity charge, the auxiliary ventilation system operating cost is: OCAuxiliary = 2,740 kW/0.70 x 365 days/year x 24 hours/day x $0.055/kWh = $1,885,900/year. Between years 2008 – 2010 and with traditional ventilation practice, the mine’s total fan power is: PTotal(2008-2010) = PPrimary = PAuxiliary = 4,459 kW + 2,740 kW = 7,199 kW. For an average 70% fan efficiency and for $0.055/kWh electricity charge, the total ventilation operating cots at “Mine A” is: OCTotal(2008-2010) = OCPrimary(2008-2010) + OCAuxiliary(2008-2010) = $3,068,920/year + $1,885,900/year = $4,954,820/year  8.2  Solving and Balancing the Primary and Auxiliary Ventilation Systems of “Mine A” for Years 2008-2010 and Activity Based Air Volume Requirements  Mining process simulations have shown that in a multi-level mining block during concurrent production and development operations (i.e. drilling, explosive loading, mucking, waste transportation) of the total number of 12 auxiliary fans assigned to the mining block, the maximum number of fans that were required to operate at any one time was 9 fans and this for limited time. Furthermore, mine process model runs showed that within the mining block, during most of the simulation time-frame, only 6 auxiliary fans were needed. As a result, with fresh air delivered when and where mining activities were performed, the total air volume required by the mining block reduced from 115 m3/s as for traditional ventilation 138  requirements to 73.0 m3/s. The air volume savings achieved within the production blocks were further transmitted to the primary ventilation system. Consequently, the mine’s total intake airflow also reduced. Figure 33 showed that during Mine A’s operating life, savings in intake air volume can vary from 168 m3/s to 252 m3/s. Remembering equation (3) from section 1.1.1, (FP = RQ3) which shows that a mine’s fan power is linearly proportional to the mine’s resistance (R) and exponentially its intake air volume, a 25% reduction in intake airflow can generated significant reduction in underground power consumption and consequently in ventilation operating cots.  8.2.1 Ventilation Modelling for Years 2008–2010 (Activity Based Ventilation) Primary System – Flow Distribution, Fan Duties and Power Requirements Mine process model runs showed that with fresh air delivered when and where mining activities were performed (activity based ventilation), the total air volume required by one mining block was: Qblock = 73.0m3/s. As in this time period ore reserves are extracted from 4 mining blocks, the total air volume required by all production activities is: Qproduction-activity = 73.0 m3/s/block x 4 mining blocks = 292 m3/s. The total air volume for secondary services and to provide minimum air volumes to the inactive areas of the mine is: Qservices = 200 m3/s. As a result, the mine’s total intake air volume becomes: QIntake-activity(2008-2010) = Qproduction-activity + Qservices = 292 m3/s + 200 m3/s = 492 m3/s. Based upon these flow requirements the primary ventilation system of “Mine A” was solved and balanced using ventilation simulation techniques. With activity based ventilation practice, the airflow distribution within the mine’s primary system is shown in Figure 36. Within the primary system, the operating characteristics of the main fans in terms of air quantity (Q) and total pressure (Ptotal) are: •  1420 Level Intake Fans (2 in parallel): Q = 135m3/s, Ptotal = 774 Pa  •  2300 Level Intake Fans (2 in parallel): Q = 220m3/s, Ptotal = 1,723 Pa  •  2600 Level Intake Fans (2 in parallel): Q = 124m3/s, Ptotal = 1,459 Pa  •  No. 11 Shaft Exhaust Fans (3 in parallel): Q = 477m3/s, Ptotal = 2,244 Pa  The operating characteristics and the operating power of all individual fans (main and boosters) that are part of the mine’s primary ventilation system are shown in Table 14.  139  Figure 36: Airflow distribution and fan operating duties in the primary system for years 2008 – 2010 - Activity based ventilation  140  Table 14: Fan operating duties and power requirements for primary system (years 2008 - 2010) – Activity based ventilation Junction From 191 39 13 380 420 500 684 487 36 701 7721 7784 7845 682 684 7724 7781 7783 7849 383 53 232 9 6 TOTAL  Junction To 190 40 29 381 421 501 702 485 35 702 7720 7783 7844 683 34 7725 7782 7790 7850 384 54 58 10 5  Pressure (kPa) 0.066 0.800 0.800 1.094 2.910 0.224 0.169 0.108 0.867 0.425 0.249 0.163 0.165 0.062 0.072 0.100 0.100 0.100 0.100 0.080 0.774 1.723 1.459 2.244  Flow Fan Power Operating Cost Fan Curve (m3/s) (kW) ($/year) Status 6.3 0.42 287 On 219.9 175.89 121,062 On 124.0 99.21 68,284 On 19.3 21.15 14,556 On 2.9 8.55 5,883 On 7.4 1.65 1,135 On 39.1 6.61 4,547 On 19.3 2.09 1,436 On 36.4 31.57 21,729 On 34.4 14.63 10,071 On 29.5 7.34 5,052 On 74.1 12.08 8,313 On 56.1 9.26 6,371 On 6.2 0.38 264 On 11.1 0.80 551 On 118.3 11.83 8,140 On 64.4 6.44 4,433 On 186.8 18.68 12,859 On 70.3 7.03 4,842 On 12.1 0.97 668 On 135.0 104.49 71,916 On 220.0 379.04 260,886 On 124.0 180.93 124,534 On 477.0 1070.38 736,725 On 2,172.00 1,494,960  Fan Arrangement 1 in Series 1 in Series 2 in Parallel 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 2 in Parallel 2 in Parallel 2 in Parallel 3 in Parallel  Fan Location 1900 LEVEL 5 SHAFT 5400L Fans 5000 Fans (2p) 3800L Booster Fan 4200L Booster Fan 5000L Booster Fan 6800-7000 R booster 4200 access booster FAR booster 7000L booster 7720L booster To 7780L booster To 7840 booster 6800 Ra drift fan 6800 - 6600 ramp FAR to ramp FAR to 7780L 7720L -7780L fan 7840-7900 FAR fan 3800L 5 Shaft 1420L Fans (2p) 2300L Fans (2p) 2600L Fans (2p) 11 Shaft Fans (3p)  141  Table 14 also shows that for years (2008–2010) and with activity based ventilation practice, the total fan power in the primary ventilation system is: PPrimary-Activity = 2,172 kW. For average 70% fans efficiency and for $0.055/kWh electricity charge, the primary ventilation system operating cost is: OCPrimary-Activity(2008-2010) = 2,172 kW/0.7 x 365 days/year x 24 hours/day x $0.055/kWh = $1,494,960/year. For the same time period but traditional ventilation practice (as determined in section 8.1.1) the total fan power in the primary system was: PPrimary-Traditional = 4,459 KW. This shows that with activity based ventilation practice, within the primary system alone, a reduction in the mine’s intake airflow of 165 m3/s (25%) can generate savings in power consumption in the range of: PSPrimary(2008-2010) = 2,287 kW/0.7 x 365 days x 24 hours/day = 28,617 MWh/year or $1,573,960/year. This represents 51% reduction in primary ventilation operating costs. Auxiliary System – Flow Distribution, Fan Duties and Power Requirements With activity based ventilation, AutoModTM model runs showed that the combined air volume required at one mining block was: QBlock = 73.0 m3/s. For these reduced airflow regimes, to determine the operating duty of the auxiliary fans and the power consumption, firstly the auxiliary ventilation system of the mining block needs to be solved and balanced. The fan operating duties, their power requirements and the airflow distribution within the mining block was determined trough ventilation simulation. During simulation exercises the conditions with respect to minimum air volume requirements along haulage levels were similar as for the auxiliary ventilation scenario with traditional ventilation practice, namely: •  Along the stope and development haulage levels: QSmin = 7 m3/s  •  Along any section of the access ramp: QRmin = 14 m3/s  Further to this, during scheduled production down-times (e.g. lunch breaks), to deliver these minimum air volumes throughout the auxiliary system and to avoid airflow recirculation at the flow pick-up location of the auxiliary fans, the combined minimum air volume required at the mining block would need to be in excess of: QBlock-min = 33.3 m3/s. With activity based ventilation practice the airflow distribution within the auxiliary ventilation system is shown in Figure 37. The operating duty of each auxiliary fan that is part of the mining block’s auxiliary ventilation system is shown in Table 15. This table also shows that the total fan power within one mining block is: PBlock-Activity = 127 kW. 142  Figure 37: Airflow distribution & fan operating duties within the mining block’s auxiliary ventilation system - Activity based ventilation  143  Table 15: Auxiliary fans operating duty and power requirements within the mining block – Activity based ventilation practice Junction  Junction  From 93 92 91 90 73 60 43 40 32 16 13 6  To 83 69 55 28 74 61 44 41 33 17 14 7  TOTAL  Pressure (kPa) 0.890 0.583 0.534 0.533 0.942 0.941 0.724 0.507 0.714 0.751 0.505 0.722  Flow  Power  (m3/s) 12.0 9.4 13.5 13.5 9.4 9.4 12.0 13.5 7.6 12.0 13.5 7.6  (kW) 14.95 7.67 10.09 10.08 12.39 12.39 12.16 9.57 7.60 12.61 9.55 7.68 127.00  Operating Fan Curve Fan Cost Status Arrangement Location ($/year) 10,291 On 1S Waste-pass #2 auxiliary fan (dev2_fan-drill) 5,281 On 1S Waste-pass #1 auxiliary fan (dev1_fan_drill) 6,947 On 1S Ore-pass #2 auxiliary fan (orepass 2) 6,934 On 1S Ore-pass #1 auxiliary fan (orepass 1) 8,532 On 1S Development #2 fan (dev2_fan_scoop) 8,523 On 1S Development #1 fan (dev1_fan_scoop) 8,372 On 1S VRM2 scavenger fan (scav_fan2_scoop) 6,595 On 1S VRM2 mucking level fan (fan2_scoop) 5,229 On 1S VRM2 drilling level fan (scav2_drill) 8,684 On 1S VRM1 scavenger fan (scav_fan1_scoop) 6,569 On 1S VRM1 mucking level fan (fan1_scoop) 5,287 On 1S VRM1 drilling level fan (scav1_drill) 87,244  144  Between years 2008 – 2010, the long-term air volume schedule (see section 7.2.2) shows that the ore reserves are extracted from 4 mining blocks. Consequently, the combined auxiliary fan power within the mining blocks is: PProduction-Activity = 127 kW/block x 4 mining blocks = 508 kW. Further to this, the mine uses 62 additional auxiliary fans to deliver fresh air for secondary services. Assuming that 40% of these auxiliary fans, namely the fans used for backfilling and ground control activities, could be fully controlled (as the production fans), the combined fan power for the secondary services would be: PServices-Activity = (24.9 kW x 37 fans) + (10.6 kW x 25 fans) = 921.3 kW + 265 kW = 1,186 kW With activity based ventilation practice the total fan power of the mine’s auxiliary ventilation system is: PAuxiliary-Activity = PProduction-Activity + PServices-Activity = 508 kW + 1,186 kW = 1,694 kW. For an average 70% auxiliary fan efficiency and for $0.055/kWh electricity charge, the operating cost of the mine’s auxiliary ventilation system is: OCAuxiliary-Activity(2008-2010) = 1,694 kW/0.7 x 365 days/year x 24 hours/day x $0.055/kWh = $1,165,960/year. This shows that for years 2008 - 2010 and with activity based ventilation, the mine’s total fan power is: PTotal-Activity(2008-2010) = PPrimary-Activity + PAuxiliary-Activity = 2,172 kW + 1,694 kW = 3,866 kW. For average 70% fan efficiency and $0.055/kWh electricity charge, the operating cost of the mine’s primary and auxiliary ventilation system becomes: OCTotal-Activity(2008-2010) = OCPrimary-Activity(2008-2010) + OCAuxiliary-Activity(2008-2010) = $1,494,960/year + $1,165,960/year = $2,660,920/year. For the same time period but traditional ventilation practice (see section 8.1.1) the mine’s total fan power and operating cost was calculated at: PTotal(2008-2010) = 7,199 kW and OCTotal(2008-2010) = $4,954,820/year, respectively. Discrete-event mining process simulations showed that for years 2008 – 2010, with activity based ventilation practice Mine A’s intake air volume could be reduced by 165 m3/s, a 25% reduction. The above calculations indicate that due to this intake air volume reduction, the combined energy consumption savings within the mine’s primary and auxiliary ventilation systems is: PSTotal(2008-2010) = (7,199 kW– 3,866 kW)/0.7 x 365 days/year x 24 hours/day = 41,710 MWh/year or $2,293,900/year, which represents 46% reduction in operating costs.  145  8.3  Power Consumption and Operating Cost Comparison – Traditional versus Activity Based Ventilation Practices  To determine the potential benefits of activity based ventilation practice, the primary and auxiliary ventilation systems of “Mine A” were solved and balanced for traditional as well as for activity based air volume requirements. For traditional ventilation practice the mine’s intake air volume was determined by summating the air volume requirements in the production areas and the air volume requirements for secondary services. Provisions in additional air volumes were also provided in order to maintain minimum airflow velocities within the inactive areas of the mine. For the production areas the required air volumes were determined based upon the diesel exhaust dilution criterion (0.063m3/kW) and were calculated according to the maximum number of diesel equipment that might operate in the production area at any one time. For activity based ventilation practice, the mine’s intake air volume was also determined by summating the air volume requirements for production, services and inactive areas. However, for the production areas the air volume requirements were determined based upon discrete-event process simulations (using AutoModTM) with fresh air volumes calculated according to diesel exhaust dilution requirements and delivered to the production workings on demand. The long-term airflow demand schedule (years 2008 – 2020) for traditional as well as for activity based ventilation took into consideration the production capacity of a typical mining block and consequently the number of mining blocks that would need to be maintained active in order to fulfill production requirements. The life-cycle airflow demand schedule for both traditional and activity based ventilation practice is shown in Figure 33. This figure also shows that during its operating life, to fulfill various production requirements the mine’s total intake air volume for traditional and activity based practices varies from 660 m3/s to 890 m3/s and from 492 m3/s to 638 m3/s, respectively. During Mine A’s operating life, to determine the potential benefits of an activity based ventilation system, its primary and auxiliary ventilation systems needed to be solved and balanced separately for the following time periods: 2008-2010, 2011-2013, 2014-2017 and 2018-2020. This was performed for traditional as well as for activity based air volume requirements. A summary with respect to Mine A’s installed fan power and associated ventilation operating costs as well as the savings generated by an activity based ventilation system versus traditional ventilation are presented in Table 16, Table 17 and Table 18, respectively, as follows: 146  Table 16: Intake Airflow, Power Consumption and Operating Costs - Traditional Ventilation Practice  No. 1 2 3 4  Time Period / Traditional Ventilation Years 2008 – 2010 Years 2011 – 2013 Years 2014 – 2017 Years 2018 – 2020  Total Intake Air Volume (m3/s) 660.0 775.0 890.0 775.0  Total Fan Power (kW) Primary 4,459 7,473 14,578 9,317  Auxiliary  Ventilation Operating Cost ($/year)  Combined  2,740 3,039 3,338 3,039  7,199 10,512 17,916 12,356  Primary  Auxiliary  3,068,920 5,143,600 10,033,830 6,413,450  1,885,900 2,091,700 2,297,500 2,091,700  Combined 4,954,820 7,235,300 12,331,330 8,505,150  Table 17: Intake Airflow, Power Consumption and Operating Costs - Activity Based Ventilation Practice  No. 1 2 3 4  Time Period / Activity Based Ventilation Years 2008 – 2010 Years 2011 – 2013 Years 2014 – 2017 Years 2018 – 2020  Total Intake Air Volume (m3/s) 492.0 565.0 638.0 565.0  Total Fan Power (kW) Primary 2,172 3,285 4,933 3,547  Auxiliary  Ventilation Operating Cost ($/year)  Combined  1,694 1,821 1,948 1,821  3,866 5,106 6,881 5,368  Primary  Auxiliary  1,494,960 2,261,020 3,395,310 2,441,350  Combined  1,165,960 1,253,370 1,340,780 1,253,370  2,660,920 3,514,390 4,736,090 3,694,720  Table 18: Power Consumption & Operating Cost Savings – Activity Based vs. Traditional Ventilation Time Period No. 1 2 3 4  Years 2008 – 2010 Years 2011 – 2013 Years 2014 – 2017 Years 2018 – 2020  Energy Consumption Savings (MWh/year) Primary Auxiliary Combined  Operating Cost Savings ($/year) Primary Auxiliary Combined  Operating Cost Savings (%) Pr. Aux. Com.  28,620.2 52,409.8 120,700.2 72,207.4  1,573,960 2,882,580 6,638,520 3,972,100  51 56 66 62  13,089.9 15,242.4 17,394.0 15,242.4  41,710.1 67,652.2 138,095.1 87,449.8  719,940 838,330 956,720 838,330  2,293,900 3,720,910 7,595,240 4,810,430  38 40 42 40  46 51 61 56  147  For the time periods of years 2008 – 2010, 2011 – 2013, 2014 – 2017 and 2018 – 2020, the intake air volume, installed fan power, ventilation operating costs and the savings in operating costs generated by an “activity based” ventilation system versus “traditional” ventilation at “Mine A” are also presented in Figure 38, Figure 39, Figure 40 and Figure 41, respectively, in a graphical format. Figure 38: Mine A’s intake air volume for “traditional” versus “activity based” ventilation Inatke Air Volume: Traditional vs. Activity Based Ventialtion 1000 890  Inatke Air Volume (m3/s)  900 775  800 700 600 500  775  660 638  565  565  492  400 300 200 100 0  2008-2010  2011-2013  2014-2017  2018-2020  Time-Period  Traditional Ventilation  Actitivy-Based Ventilation  148  Figure 39: Mine A’s installed fan power for “traditional” versus “activity based” ventilation Installed Fan Power: Traditional vs. Activity Based Ventialtion  17,916  18,000 Installed Fan Power (kW)  16,000 14,000  12,356  12,000  10,512  10,000 8,000  6,881  7,199  4,000  5,368  5,106  6,000 3,866  2,000 0  2008-2010  2011-2013  2014-2017  2018-2020  Time-Period  Traditional Ventilation  Actitivy-Based Ventilation  Figure 40: Mine A’s operating costs for “traditional” versus “activity based” ventilation Ventilation Operting Cost: Traditional vs. Activity Based Ventialtion $14.00 Operating Cost ($M/year)  $12.33M $12.00 $10.00  $8.50M $7.23M  $8.00 $6.00 $4.00  $4.95M  $4.73M $3.51M  $3.69M  $2.66M  $2.00 $0.00  2008-2010  2011-2013  2014-2017  2018-2020  Time-Period  Traditional Ventilation  Actitivy-Based Ventilation  149  Figure 41: Savings in operating costs generated by an “activity based” ventilation system versus “traditional” ventilation at “Mine A” Savings in Operating Cost ($M/Year)  Operating Cost Savings ($M/Year)  8  $7.60M (61% ) $0.96M  7 6  $4.81M (56% )  5  3  $6.64M  $0.84M $2.29M (46% )  2  $0.72M  1  $1.57M  0  $0.84M  $3.72M (51% )  4  2008-2010  $3.97M  $2.88M  2011-2013  2014-2017  2018-2020  Time-Period  Primary System  Auxilary System  Figure 41 shows that the magnitude of operating cost savings generated by an activity based ventilation system at “Mine A” are more significant within its primary ventilation system comparatively to its auxiliary ventilation system. This clearly shows that in order to achieve the full benefits of an activity based ventilation system, the reductions in intake airflow that are achieved within the mine’s auxiliary ventilation system would need to be transmitted to the primary ventilation system, namely to the main surface and booster fans.  150  9.  ECONOMIC AND ENVIRONMENTAL BENEFITS OF THIS NEW VENTILATION DESIGN APPROACH  Mining process simulations performed on a typical mining block that employs vertical retreat mining have shown that an activity based auxiliary ventilation system where appropriate air volumes are delivered to the production and development workings on-demand, in other words when, where and only as long as mining activities are performed, can generate savings in intake air volume requirements for the mining block in the range of 37%. These savings are comparable to savings in intake airflow generated by activity based ventilation control systems reported by other underground operations world-wide. For example, at LKAB’s Malmberget mine, within its mining blocks, the saving in intake airflow generated by and event based ventilation control system are in the range of 39% [Nensen & Ludkvinst, 2005]. Furthermore, at Creighton mine’s 2340m level, the saving in intake airflow generated by an event based ventilation control system is reported at 40% [O’Connor, 2008]. Calculations showed that when an activity based auxiliary ventilation system was applied to an existing deep metal mine (Mine A), the savings achieved within its production blocks were further transmitted to the mine’s primary ventilation system resulting in a reduction in the mine’s total intake airflow. The life-cycle airflow demand schedule carried out in section 7.2 (see Figure 33), shows that with an activity based auxiliary ventilation system, the savings in Mine A’s intake air volume vary from 168 m3/s (25%) to 252 m3/s (28%), as a function of the mine’s production requirements and the configuration of its primary ventilation infrastructure.  9.1  Cost Estimate of an Event-Based Ventilation Control System  With this level of control, the air volume requirements and usage within a multi-level metal mine’s auxiliary ventilation system are determined automatically from information provided by the mine’s vehicle identification, tracking and tagging system. The electronic tag board is still required to provide manual means of declaring ventilation requirements for mining equipment (i.e. personnel carriers) that may not be equipped with tracking and tagging systems and to assign air volumes to the inactive areas of the mine. The air volume requirements within the mine’s primary ventilation system are also determined automatically from information gathered from the auxiliary ventilation system and ventilation parameters entered through the electronic tag board. 151  The ventilation control system architecture for each production level comprises the following equipment [Bestech Engineering, 2007]: •  Variable airflow regulator at the bottom of the intake raise through which fresh air is delivered to the production and development workings  •  Variable frequency drive on the auxiliary fans assigned to production stope, development and haulage levels  •  Vehicle tracking and tagging system  •  Two air flow transmitters to monitor the air volume entering the production stope (via the auxiliary ducting system) and the air volume exhausted from the production stope along the haulage drift  •  Two air quality transmitters able to monitor CO, CO2, NO, NO2, one located on the stope haulage drift close to the RAR and one located in the development heading.  •  One air temperature transmitter located near the level’s return air raise  •  Electronic tag board, containing indicator lights (green light – permission to enter, yellow light – caution, red light – do not enter, etc.) and pushbuttons to allow mobile equipment operator whose vehicle is not equipped with electronic tags to identify the mining equipment and provide additional airflow usage.  •  Local PLC to control and monitor all auxiliary fans, electronic tag board that interacts with NRG-1  •  Ethernet control network extension to the PLC  In addition to the production levels the ventilation control system includes the following equipment: •  Variable frequency drive on the main fans (already installed at Mine A)  •  Local PLCs to control the main intake and exhaust fans of the mine’s primary ventilation system  •  NRG-1 web-based Energy Management System located on surface  The budgetary cost estimate of a fully automated ventilation control system per production level includes [Bestech Engineering, 2007]: • •  Variable frequency drives for the auxiliary fans assigned to the production stope (5 units - mucking, drilling, development, ore-pass, waste-pass):  $75,000  Vehicle identification and tracking system (estimate):  $50,000 152  •  Automated regulator at the bottom if the intake raise (1):  $35,000  •  PLC panel and hardware (1):  $25,000  •  Electronic tag board (1):  $15,000  •  Air quality instruments on the level (5):  $10,000  •  Air quality instruments on production & development LHDs (2):  $5,000  •  Ethernet network extension to the level:  $15,000  •  Construction material and labour:  $75,000  •  Engineering services:  $20,000  Per Level Total  $325,000  In addition to the production levels the estimated costs of the automated ventilation control system include the following surface equipment [Bestech Engineering, 2007]: •  NRG-1 energy management system:  $50,000  •  NRG-1 software programming and engineering services:  $100,000  Surface Total  $150,000  During its operating life the distribution of the capital expenditure of the automated ventilation control system is presented below: •  Year 2009: installation of the NRG-1 energy management system and automated control systems on eight production levels (between 7200L-7900L): Total Cost2009 = $150,000/surface equipment + $325,000/level x 8 levels = $2,750,000  •  Year 2011: installation of ventilation control equipment on three additional production levels (7960L, 8020L & 8080L): Total Cost2011 = $325,000/level x 3 levels = $975,000  •  Year 2013: installation of ventilation control equipment on two additional production levels (8140L & 8200L): Total Cost2013 = $325,000/level x 2 levels = $650,000  9.2  Cost Estimate for a New Intake System at “Mine A”  The long-term planning data shows that between years 2014 – 2017 the production requirements at “Mine A” will increase from 5,000 to 6,000 tons/day. To fulfill this requirement, the mine would need to maintain 6 mining blocks (or 12 production stopes) active. Consequently, with traditional ventilation practice the mine’s total intake air volume will increase from 775 m3/s to 890 m3/s. 153  Computer simulations performed on Mine A’s ventilation model (years 2014 – 2017 configuration), showed that this air volume requirement is well beyond the capacity of the current ventilation infrastructure. To deliver this amount of airflow through its primary system, eight intake fans would need to be replaced with larger fans and more powerful motors. Preliminary calculations showed that the associated capital and ventilation operating costs in regard to the replacement of the intake fans could increase to an unsustainable level. One solution proposed by the mine’s engineering department was to reduce the overall resistance of the mine and consequently the pressure requirements from the intake fans by constructing a new intake fresh air system as follows: •  Connect No. 5 Shaft and No. 8 internal shaft with a 4.2m diameter raisebore driven from the 3800L to the 5000L  •  Eliminate all level fans that were used to direct return air into the No. 5 Shaft  •  Install ventilation doors on all access drifts that connect the upper levels to No. 5 Shaft  •  Install two new intake fans (type 10150-AMF-6600) on the top of the No. 5 shaft in parallel. This will reverse the airflow along the No. 5 Shaft  With the new intake system and traditional ventilation practice the mine’s primary ventilation system was solved and balanced for years 2014-2017 and years 2018-2020 configurations. For these time periods, the airflow distribution within the mine’s primary ventilation system is shown in Figure 42 and Figure 43, respectively.  154  Figure 42: Airflow distribution in the primary ventilation system with new intake system for years 2014 - 2017 - Traditional ventilation  155  Figure 43: Airflow distribution in the primary ventilation system with new intake system for years 2018 - 2020 - Traditional ventilation  156  Between years 2014-2017, with the new intake system and traditional ventilation, the operating duty and the fan power of all primary fans is shown in Table 19. This table also shows that within this time period and as a result of this new intake infrastructure, the total fan power of the mine’s primary system reduces from 14,578 kW (as calculated in section B.1.2) to 11,890 kW. As the auxiliary fan power remains the same (as calculated in section B.1.2), between years 2014-2017 and with this new intake infrastructure the mine’s total fan power becomes: PNewIntake(2014-2017)  = 11,890 kW + 3,338 kW = 15,228 kW. For average 70% fans efficiency and  $0.055/kWh electricity charge, the total ventilation operating cost is: OCNew-Intake(2014-2017) = 15,228 kW/0.7 x 365 days x 24 hours/day x $0.055/kWh = $10,481,210/year. Between years 2018-2020, with the new intake system and traditional ventilation practice the operating duty and the fan power of all primary fans is shown in Table 20. Within this time period and as a result of this new intake infrastructure, the total fan power of the mine’s primary system reduces from 9,317 kW (as calculated in section B.1.3) to 7,668 kW. Again, as the auxiliary fan power remains the same (as calculated in B.1.3), between years 2018-2020 and with this new intake infrastructure the mine’s total fan power becomes: PNewIntake(2018-2020)  = 7,668 kW + 3,090 kW = 10,758 kW. For average 70% fans efficiency and  $0.055/kWh electricity charge, the total ventilation operating cost is: OCNew-Intake(2018-2020) = 10,758 kW/0.7 x 365 days/year x 24 hours/day x $0.055/kWh = $7,404,580/year  157  Table 19: Fan power in the primary system with new intake infrastructure (Years 2014 - 2017) – Traditional ventilation practice Junction From 7726 7788 7853 7908 7969 8090 8152 8210 8032 8205 8208 13 39 487 36 7844 7851 7906 7965 8025 8146 8083 53 6 232 9 38 42 TOTAL  Junction To 7727 7789 7854 7909 7970 8091 8153 8211 8033 8206 8209 29 40 485 35 7843 7852 7907 7966 8026 8147 8084 54 5 58 10 41 43  Pressure (kPa) 0.669 0.684 0.311 0.313 0.346 0.413 0.452 0.544 0.365 0.26 0.27 0.5 3.5 0.127 1.973 0.431 0.217 0.044 0.085 0.121 0.198 0.16 3.8 8.325 3.818 1.219 0.819 0.819  Flow Fan Power Operating Cost (m3/s) (kW) ($/year) 15.19 10.16 6,995 26.16 17.89 12,314 106.95 33.26 22,893 47.92 15.00 10,323 2.34 0.81 558 63.21 26.10 17,967 50.82 22.97 15,812 56.81 30.90 21,271 15.2 5.55 3,818 27.66 7.19 4,949 15.42 4.16 2,865 133.03 66.51 45,781 448.02 1568.06 1,079,270 19.46 2.47 1,701 39.79 78.50 54,030 28.52 12.29 8,459 26.34 5.72 3,934 96.74 4.26 2,930 46.98 3.99 2,749 20.01 2.42 1,666 10.93 2.16 1,490 11.67 1.87 1,285 145.25 551.94 379,891 890.3 7411.74 5,101,396 448.01 1710.50 1,177,313 133.03 162.16 111,614 79.98 65.51 45,086 79.98 65.51 45,086 11,890.00 8,183,446  Fan Curve Status On On On On On On On On On On On On On On On On On On On On On On On On On On On On  Fan Arrangement 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 2 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 2 in Parallel 3 in Parallel 2 in Parallel 2 in Parallel 1 in Parallel 1 in Parallel  Fan Location 7720L Booster Fan 7780L Booster Fan 7840L Booster Fan 7900L Booster Fan 7960L Booster Fan 8080L Booster Fan 8140L Booster Fan 8200L Booster Fan 8020L Booster Fan 8200L booster 8200L booster fan 5000L Fans (2p) 5400L Fans 6 shaft x-cut on 4800L FAR 7200-7400 7840L fan 7840L to FAR/Esc. 7900L fan 7960L fan 8020L fan 8140L to ramp 8080L fan 1420L Fans (2p) 11 Shaft fans (3p) 2300L Fans (2p) 2600L Fans (2p) New surface fans New surface fans  158  Table 20: Fan power in the primary system with new intake infrastructure (years 2018 - 2020) – Traditional ventilation practice Junction From 7726 7788 7853 7908 7969 8090 8152 8210 8032 8205 8208 13 39 487 36 7844 7851 7906 7965 8025 8146 8083 53 6 9 38 42 232 TOTAL  Junction To 7727 7789 7854 7909 7970 8091 8153 8211 8033 8206 8209 29 40 485 35 7843 7852 7907 7966 8026 8147 8084 54 5 10 41 43 58  Pressure (kPa) 0.669 0.684 0.311 0.313 0.346 0.413 0.452 0.544 0.365 0.260 0.270 0.250 2.000 0.127 1.973 0.431 0.217 0.044 0.085 0.121 0.198 0.160 2.944 6.342 0.767 1.132 1.132 3.004  Flow Fan Power Operating Cost Fan Curve (m3/s) (kW) ($/year) Status 16.30 10.90 7,505 On 22.40 15.32 10,547 On 103.84 32.3 22,228 On 43.18 13.52 9,303 On 0.30 0.11 72 On 63.07 26.05 17,929 On 50.74 22.93 15,785 On 56.80 30.90 21,266 On 14.25 5.20 3,579 On 27.65 7.19 4,949 On 15.42 4.16 2,865 On 147.78 36.95 25,429 On 368.86 737.73 507,769 On 19.79 2.51 1,730 On 4.83 9.53 6,563 On 37.72 16.26 11,191 On 26.89 5.83 4,016 On 96.68 4.25 2,928 On 47.90 4.07 2,802 On 19.86 2.40 1,654 On 10.94 2.17 1,490 On 11.67 1.87 1,285 On 143.11 421.32 289,989 On 773.88 4907.98 3,378,091 On 147.78 113.35 78,016 On 54.95 62.20 42,815 On 54.95 62.20 42,815 On 368.99 1108.45 762,930 On 7,668.00 5,277,541  Fan Arrangement 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 2 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 2 in Parallel 3 in Parallel 2 in Parallel 1 in Parallel 1 in Parallel 2 in Parallel  Fan Location 7720L Booster Fan 7780L Booster Fan 7840L Booster Fan 7900L Booster Fan 7960L Booster Fan 8080L Booster Fan 8140L Booster Fan 8200L Booster Fan 8020L Booster Fan 8200L booster 8200L booster fan 5000L Fans (2p) 5400L Fans 6 shaft x-cut on 4800 L FAR 7200-7400 7840L fan 7840L to FAR/Esc. 7900L fan 7960L fan 8020L fan 8140L to ramp 8080L fan 1420L fans (2p) 11 Shaft Fans (3p) 2600L fans (2p) New surface fans New surface fans 2300L Fans (2p)  159  The cost estimate of Mine A’s new intake infrastructure includes a 4.2m diameter raisbore that connects the No. 5 Shaft to the No. 8 internal shaft and the following surface equipment [HATCH Engineering, 2008]: •  Two 10150-AMF-660 surface fans installed on top of the No. 5 Shaft in parallel arrangement: 2 x $306,250/fan  = $612,500  •  Two 260 kW (350hp) electric motors: 2 x $85,000/motor  = $170,000  •  Surface heater house plus natural gas burners  = $890,000  •  Fan silencers: 2 x $50,000/silencer  = $100,000  •  Raise bore (4.2m diameter) connecting No. 5 and No. 8 Shafts: 366m raise x $3,230/m  •  Engineering services: New Intake Infrastructure Totals  = $1,182,180 = $200,000 = $3,154,680  For the operating life of “Mine A”, the combined ventilation operating costs (primary + auxiliary systems) determined in section 8 (years 2008-2010), and Appendix B (years 20112013, 2014-2017, 2018-2020), the associated capital investments and their distribution for traditional ventilation practice (without and with a new intake system) versus activity based ventilation practice are shown in Table 21 and Table 22, as follows:  160  Table 21: Ventilation Operating & Capital Costs for Activity Based vs. Traditional Ventilation – Without New Intake Infrastructure TRADITIONAL VENTILATION PRACTICE ACTIVITY BASED VENTILATION PRACTICE Without New Intake Infrastructure Years Ventilation Capital Costs Capital Total Costs Potential Capital Total Costs Ventilation Capital Costs for New Costs for Operating Ventilation Operating for New Costs for Operating Operating Savings Cost Ventilation Automated + Cost Ventilation Automated + Infrastructure Control Capital Infrastructure Control Capital ($/year) ($/year) System System ($/year) ($/year) ($/year) ($/year) ($/year) ($/year) ($/year) 2008 4,954,820 4,954,820 4,954,820 4,954,820 0 2009 4,954,820 4,954,820 2,660,920 5,410,920 -456,100 2,750,000 8 levels 2010 4,954,820 4,954,820 2,660,920 2,660,920 2,293,900 2011 7,235,300 7,235,300 3,514,390 4,489,390 2,745,910 975,000 3 levels 2012 7,235,300 7,235,300 3,514,390 3,514,390 3,720,910 2013 7,235,300 7,235,300 3,514,390 4,164,390 3,070,910 650,000 2 levels 2014 12,331,330 12,331,330 4,736,090 4,736,090 7,595240 2015 12,331,330 12,331,330 4,736,090 4,736,090 7,595240 2016 12,331,330 12,331,330 4,736,090 4,736,090 7,595240 2017 12,331,330 12,331,330 4,736,090 4,736,090 7,595240 2018 8,505,150 8,505,150 3,694,720 3,694,720 4,810,430 2019 8,505,150 8,505,150 3,694,720 3,694,720 4,810,430 2020 8,505,150 8,505,150 3,694,720 3,694,720 4,810,430 Totals 111,411,130 4,375,000 55,223,350 56,187,780  Table 22: Ventilation Operating & Capital Costs for Activity Based vs. Traditional Ventilation – With New Intake Infrastructure  161  TRADITIONAL VENTILATION PRACTICE With New Intake Infrastructure Years  Ventilation Operating Cost  2008 2009  ($/year) 4,954,820 4,954,820  2010 2011  4,954,820 7,235,300  2012  7,235,300  2013  7,235,300  2014 2015 2016 2017 2018 2019 2020 Totals  10,481,210 10,481,210 10,481,210 10,481,210 7,404,580 7,404,580 7,404,580  ACTIVITY BASED VENTILATION PRACTICE  Potential Capital Total Costs Ventilation Capital Costs Capital Total Ventilation Costs for Operating Operating for New Costs for Costs Savings Automated + Cost Ventilation Automated Operating Control Capital Infrastructure Control + System System Capital ($/year) ($/year) ($/year) ($/year) ($/year) ($/year) ($/year) ($/year) 4,954,820 4,954,820 4,954,820 0 4,954,820 2,660,920 -456,100 2,750,000 5,410,920 8 levels 4,954,820 2,660,920 2,660,920 2,293,900 7,235,300 3,514,390 2,745,910 975,000 4,489,390 3 levels 8,417,480 3,514,390 3,514,390 4,903,090 1,182,180 raise constr. 9,207,800 3,514,390 5,043,410 1,972,500 650,000 4,164,390 surface equip. 2 levels 10,481,210 4,736,090 4,736,090 5,745,120 10,481,210 4,736,090 4,736,090 5,745,120 10,481,210 4,736,090 4,736,090 5,745,120 10,481,210 4,736,090 4,736,090 5,745,120 7,404,580 3,694,720 3,694,720 3,709,860 7,404,580 3,694,720 3,694,720 3,709,860 7,404,580 3,694,720 3,694,720 3,709,860 3,154,680 103,863,620 4,375,000 55,223,350 48,640,270  Capital Costs for New Ventilation Infrastructure  162  Table 21 shows that with traditional ventilation practice and without an additional intake infrastructure the combined ventilation operating plus capital costs at “Mine A” from year 2008 to year 2020 are in the range of $111.4M, which represents average ventilation operating costs of $8.57M/year. Furthermore, during the same time period but with activity based ventilation practice, the combined ventilation operating plus capital costs are in the range of $55.2M, which represents average ventilation capital and operating costs of $4.27M/year. Table 21 also shows that the combined capital investments of a fully automated ventilation control system at “Mine A” is $4,375,000, of which $2,750,000 would be used in 2009 to install the components of the ventilation controls system (as described in section 9.1) on eight production levels. Furthermore, $975,000 and $650,000 would be used in 2011 and 2013, respectively to install ventilation control equipment on five additional production levels as production requirements at “Mine A” increase. The last column in Table 21 shows the potential ventilation savings on a yearly basis as well as for the entire operating life of the mine. This column shows that the potential ventilation savings in 2009 have a negative value. This is due to the fact that nearly 63% of the ventilation control system’s capital investment ($2,750,000) would be used in 2009. However, in the following year (2010), the potential savings of the activity based ventilation practice are in the range of $2.29M. This is because 83% (or $2,293,000) of the capital investment that was used in 2009 will already be recovered by the operating cost savings generated by the activity based ventilation control system. During its operating life and with no additional intake infrastructure, the potential ventilation savings at “Mine A” would be in the range of $56.1M, which on average represents $4.31M/year (50.4%) reduction in operating costs. Table 22 shows that with traditional ventilation practice and an additional intake infrastructure the combined ventilation operating plus capital costs at “Mine A” from 2008 to 2020 would be in the range of $103.8M. In this scenario, the combined ventilation costs are approximately $7.6M lower than the combined ventilation costs in the previous scenario where the mine would operate without an additional intake system. In other words, during its operating life and with traditional ventilation practice, a $3.15M capital investment needed to construct a new intake system would reduce the mine’s operating cost by $7.5M This represents average operating cost reductions of $0.57M/year. Furthermore, the last column in 163  Table 22 shows that with activity based ventilation practice the combined ventilation savings at “Mine A” between 2008 and 2020 would be in the range of $48.6M, which on average represents $3.7M/year reduction in operating costs. This basically shows that the potential cost savings that would be generated by an activity based ventilation system ($3.7M/year) are significantly higher than the cost savings that would generated by a new intake system ($0.57M/year). This can be explained by referencing equation 3 (FP = RQ3) which states that the mine’s fan power and consequently, its ventilation operating cost, is linearly proportional to the equivalent resistance of the mine (R) and proportional to the 3rd power of the mine’s intake airflow. The limited benefits of a new intake infrastructure can also be observed when comparing Figure 42 to Figure 46, which display the fan operating duties and the airflow distribution throughout the mine’s primary ventilation system between years 2014 – 2017 for both scenarios, with and without the additional intake infrastructure. Figure 42 shows that the new intake system would reduce the mine’s equivalent resistance and the new surface fans installed on the top of the new intake system would also reduce the operating pressures of the main fans located on the 2300L and the 2600L. However, ventilation simulations show that the reduction in the mine’s equivalent resistance is not significant and the new surface fans would not reduce the operating pressure on the main exhaust fans located on top of the No. 11 Exhaust Shaft. This concludes that at “Mine A”, reducing the mine’s total intake air volume by means of an activity based automated ventilation control system would be far more beneficial than any attempt to reduce the mine’s equivalent resistance through additional new intake/exhaust infrastructures and additional pressure generators. Table 22 presented and discussed above, is also presented in Figure 44 in a graphical format.  164  Figure 44: Combined capital and operating costs for “traditional” versus “activity based” ventilation at “Mine A”  Ventilation Costs: Traditional vs. Activity-Based ($/Year) 12,000,000  Traditional Ventilation  $2.0M Capital Investment  10,000,000  10.4M  9.2M 8.4M  $2.7M Capital Inv.  Total Cost ($/Year)  10.4M 10.4M 10.4M  $1.1M Capital Investment  8,000,000  7.4M 7.4M 7.4M  7.2M  Average Ventilation Savings = $3.7M/Year  6,000,000  4.9M 4.9M  4,000,000  5.4M  4.9M  4.9M  4.7M 4.7M  4.5M  4.7M  4.7M  4.2M  3.7M  3.5M  3.7M  3.7M  2.6M  2,000,000  Activity-Based Ventilation $1.0M Capital Investment  0 2008  2009  2010  2011  $0.6M Capital Investment  2012  2013  2014  2015  2016  2017  2018  2019  2020  Years  Total Cost-Traditional  Total Cost-Activity Based  165  9.3  Environmental Benefits of this New Ventilation Design Approach  9.3.1 Green House Gas Emissions – Background Information According to the Canada’s emissions outlook, the industrial sector, which includes all manufacturing industries, namely the forestry, mining and construction sectors (excludes the oil & gas sector), is the largest energy-using sector. In 1997, the industrial sector produced 127 million tonnes (Mt) of equivalent CO2 emissions in greenhouse gases. Two-thirds of these emissions were generated by the consumption of energy and the remaining one third from various industrial processes such as CO2 release from cement production and anode production in aluminum smelting and sulphur hexafluoride use in magnesium smelting. The six energy intensive industries such as chemicals, petroleum refining, mining and smelting, iron & steel, pulp & paper, and cement account for over 80% of the industrial sectors emissions [Canada’s Emissions Outlook, 1999], [Canada’s Energy Outlook, 1998]. With regard to GHG emissions by sector, the fossil fuel industry is the largest contributor to emissions growth, with increases of approximately 64% between 1990 and 2010. This largely reflects the increase in oil sands production anticipated to occur during this period. The increase in GHG emissions from the industrial sector is also significant but the pace is somewhat slower, primarily due to greater energy efficiency improvements and reductions in certain process emissions (i.e. SF6 from magnesium smelting). For electricity generation, emissions grow rapidly from 1990 to 2015. After 2015, however, growth declines sharply as existing coal-fired plants, reaching the end of their service life, are retired and replaced by natural gas or highly-efficient coal-fired plants [Kocsis et al, 2003]. Table 23 shows the emissions anticipated from the industrial sector from various industries and groupings. Over the 2000-2020 time period, total industrial emissions are projected to grow from 125 Mt to 152 Mt. This average growth is relatively slow (1% per year), and is generally due to a shift to less energy intensive industries and efficiency gains. While the emissions increase for the sector is approximately 10%, not all industrial sub-sectors are experiencing increases. Among the industry sub-sectors, the largest increases compared to 1990 emissions level are projected for construction (71%), metal mining (64%), petroleum refining (31%) and cement (24%) [Canada’s Emissions Outlook, 1999].  166  Table 23: GHG Emissions in CO2 Equivalent (Mt) by Industry Sub-Sectors [Canada’s Emissions Outlook, 1999] Industry Sub-Sectors Pulp and Paper Chemicals Iron and Steel Smelting and Refining Petroleum Refining Other Manufacturing Metal Mining Construction Cement Forestry TOTAL  1990 (Mt) 12.6 26.8 14.1 14.1 17.8 24.3 4.7 0.7 9.7 1.0 125  2000 (Mt) 11.3 20.2 16.0 12.8 21.4 24.6 6.9 1.3 10.0 0.7 125  2010 (Mt) 13.3 23.0 16.2 12.6 23.3 27.2 7.7 1.2 12.0 0.7 138  2020 (Mt) 13.7 26.9 17.5 13.2 26.9 29.3 8.5 1.2 14.7 0.7 152  9.3.2 Projected Reductions in GHG Emissions at “Mine A” Generated by an Activity Based Ventilation Control System Within the mining industry it is a common practice that the environmental as opposed to potential monetary benefits (i.e. cost savings) that are generated by a new design approach and/or implementation of new technologies are overlooked. Table 23 (section 9.3.1), shows that the underground operations through their use of electricity and fossil fuels for intake airflow heating in winter will be responsible for generating approximately 7.7 million tonnes of equivalent CO2 emissions by 2010 and 8.5 million tonnes of equivalent CO2 emissions by 2020. On the other hand, current emissions data indicate that in order to comply with the “Kyoto Protocol” the GHG emissions generated by the Canadian industrial sector would need to be significantly reduced. Consequently, the Canadian industrial sector, including the mining industry would need to find viable solutions in order to reduce their GHG emissions without compromising production output. In this regard, underground ventilation systems designed according to “true” ventilation requirements with primary and auxiliary airflows managed through the use of an activity based ventilation control system can be extremely beneficial in reducing the mining industry’s power consumption and consequently its GHG emissions. For example, assuming that with traditional ventilation, to reduce the resistance of the mine and consequently the primary fans operating pressures, “Mine A” would design the primary ventilation system to include the additional intake infrastructure (as described in section 9.2) 167  by year 2014. For this scenario, the combined ventilation operating costs for traditional and activity based ventilation practices are shown in Table 24. The last column in Table 24 shows the potential reduction in GHG emissions associated to the reductions in underground energy consumption expressed in tonnes of equivalent CO2 per year. In order to express the reductions in GHG emissions in equivalent CO2 a conversion factor of 200 tonnes of equivalent CO2 emissions per giga watt hour (GWh) was used. This conversion factor is specific to Canada in that it reflects the make-up of the country’s electricity generation, namely from coal-fired, hydro and nuclear power plants. Table 24: Reductions in GHG emissions generated by ventilation control system at “Mine A” Ventilation Operating Costs Year  2008 2009 2010 2011 2012 2013 2014 2015 2016 2017 2018 2019 2020 Totals  Traditional Ventilation Practice  Activity Based Ventilation Practice  ($/year)  ($/year)  4,954,820 4,954,820 4,954,820 7,235,300 7,235,300 7,235,300 10,481,210 10,481,210 10,481,210 10,481,210 7,404,580 7,404,580 7,404,580 100,708,940  4,954,820 2,660,920 2,660,920 3,514,390 3,514,390 3,514,390 4,736,090 4,736,090 4,736,090 4,736,090 3,694,720 3,694,720 3,694,720 50,848,350  SAVINGS/REDUCTIONS Activity Based vs. Traditional Ventilation Operating Energy Equivalent Costs Consumption CO2 Emissions ($/year)  0 2,293,900 2,293,900 3,720,910 3,720,910 3,720,910 5,745,120 5,745,120 5,745,120 5,745,120 3,709,860 3,709,860 3,709,860 49,860,590  (MWh/year)  0.0 41,707.3 41,707.3 67,652.9 67,652.9 67,652.9 104,456.7 104,456.7 104,456.7 104,456.7 67,452.0 67,452.0 67,452.0 906,556.2  (tonnes/year)  0.0 8,341.5 8,341.5 13,530.6 13,530.6 13,530.6 20,891.3 20,891.3 20,891.3 20,891.3 13,490.4 13,490.4 13,490.4 181,311.2  168  10.  THESIS SUMMARY – RESULTS AND DISCUSSIONS  In recent years, research studies have shown that in underground operations job satisfaction, safety and workers’ productivity closely correlate with the quality of the underground environment [Schutte et al, 1997, 2007] [Payne et al, 2008]. Consequently, the environmental standards within the underground operations have been raised substantially. Despite the fact that in the last decade technological advances within the mining industry significantly reduced labour intensiveness, the need to maintain good working conditions underground is still challenging. This is due to the fact that as Canadian mines deepen ventilation and ground control related costs could eventually climb to unsustainable levels. Furthermore, forecasts of increased energy prices and new commitments with regard to reductions in GHG emissions have determined the Canadian mines to look into developing new concepts and design tools to lower production costs while meeting commitments with respect to reducing energy consumption underground. To identify the underground operations that would benefit the most from this new ventilation design approach, based upon “life-cycle” airflow demand schedules determined by means of mining process simulation, an initial assessment of all mining methods employed by Canadian mines was performed. This assessment indicated that one of the most important elements that can influence the design of a mine’s primary and auxiliary ventilation systems is the employed mining method(s). This is because the selected mining method can ultimately dictate the layout of the development and production stopes, the ground support method and the ore/waste transportation systems, and thus the size and number of production and service equipment. When optimizing a mine’s ventilation system, the first and most important step is to determine its “true” intake air volume requirement, in other words to identify and determine the ventilation redundancy in the ventilation system. In this regard, the two case studies presented in section 5, demonstrated that the magnitude of ventilation redundancy can be used to determine the efficiency of complex underground ventilation systems. The first case study showed that determining the ventilation redundancy within the mine’s auxiliary ventilation system mine-wide was problematic due to a large number of auxiliary fans and the lack of data as to when and for how long these fans were operating. Furthermore, during the 125 days of the monitoring time-frame, mine activity logs showed 169  that on average the auxiliary fans located in a production block that employed vertical retreat mining as the extraction method were required to operate at their maximum capacity only 20% of the available shift time. The second case study showed again that determining ventilation redundancy in a large and heavily mechanized base-metal mine is not straightforward due to the dynamic nature of the production operations continually redefining where ventilation is required to provide adequate working conditions underground. In this study, the mine’s intake air volume was determined using two airflow evaluation methods. Both methods were shown to have limitations in determining the mine’s intake air volume requirement: •  The future “production planning” based analysis (2nd stage review) underestimated the mine’s total intake air volume due to the fact that no considerations were given to prevent uncontrolled recirculation and provide minimum air volumes along the nonproductive levels.  •  The “SIMS database” analysis (3rd stage review) captured the dynamic nature of the development and production operations. However, this method appeared to overestimate the mine’s intake air volume due to the fact that the diesel equipment operating in the production workings did not have a “time stamp”. Consequently, fresh air was allocated multiple times to the same equipment reporting to a specific production level but at different times.  The activity based air volume requirement of a multi-level mining block was initially evaluated through discrete-event process simulation trials performed on an early version of the “Mining Process” model. These process simulation trials identified a number of model development deficiencies as described in section 6.2.4. In addition, mining activity simulation trials performed on this early version of the process model showed that the operations within various production cycles did not exactly match the operation sequence observed at an existing mine that employs the VRM method. To correct these model development deficiencies, the structure of the process logic that controls all mining and fan activities within the “Mining Process” model as well as its data entry utility were redesigned. Furthermore, within the process system new subroutines for the drilling, explosive loading, basting and ore/waste transportation were developed in order to synchronize the sequence of the mining operations simulated by the “Mining Process” model with the operating sequence observed at an existing mine. Ventilation output data generated through discrete-event 170  process simulation performed on the redesigned and improved version of the “Mining Process” model showed that the flexibility and functionality of its process system has the following capabilities: •  The fan operating duties could be entered or easily modified through the use of an MSExcelTM spreadsheet utility, namely the “Fan Volume” worksheet.  •  For each particular fan assigned to the mining block, the “Fan Volume” worksheet enables the user to enter “early fan start” times. This function is important in cases where the auxiliary fans delivering airflow to the draw-points, needed to be turned “On” before the mining equipment enters the production area.  •  To avoid uncontrolled airflow recirculation, the “Fan Summary” macro has the ability to automatically assign an additional 25% of air volume to the auxiliary fans that deliver fresh air to another independent auxiliary system.  •  When determining the activity based intake air volume requirement of a mining block, the process system in conjunction with its variables and model initiation functions has the capability of taking into account scheduled and unscheduled production down-times (e.g. lunch breaks, shift changes, blasting fume clearance, equipment failure)  From the “Custom Report” data file that was generated through AutoModTM simulations performed on the “Mining Process” model, the “Fan Summary” macro had the ability to determine the mining block’s intake air volume requirement. With the “early fans start” condition imposed on all auxiliary fans that delivered airflow for operations that immediately followed stope/development blasting fumes clearance times, the “Fan Summary” macro calculated that on average, the combined activity based air volume required by the mining block was Qactivity-block = 73.0 m3/s. For the same mining block but with traditional ventilation practice the combined air volume required at the mining block was: Qtraditional-block = 114.7 m3/s. This showed that an activity based ventilation control system could reduce the intake air volume requirements of this mining block by 41.7 m3/s, a 36% reduction. The savings in intake airflow generated within the mining blocks were then further transmitted to the mine’s primary system as well. For practical reasons and to accurately determine the potential capital and operating costs benefits of an activity based ventilation control system, this new mine ventilation design concept was applied to an existing large and deep base-metal mine (Mine A). Based upon the 171  number of mining blocks that are needed to achieve the production targets and their air volume requirements, Mine A’s life-cycle airflow demand schedule was determined for traditional and activity based ventilation practices (see section 7.2). Mine A’s intake air volume requirements and the saving in intake airflow generated by an activity based ventilation system are shown in Table 25, as follows: Table 25: Mine A’s life-cycle airflow demands for traditional & activity based ventilation  Time Period  Traditional Ventilation (m3/s)  Activity-Based Ventilation (m3/s)  Intake Airflow Savings (m3/s)  Intake Airflow Savings (%)  Years 2008-2010 Years 2011-2013 Years 2014-2017 Years 2018-2020  660 775 890 775  492 565 638 565  168 210 252 210  25 27 28 27  For theses time periods and associated intake air volume requirements, Mine A’s primary and auxiliary ventilation systems were solved and balanced using VnetPCTM ventilation program. This allowed the determination of the mine fan operating duties, their power requirements and the ventilation system operating costs. For Mine A’s operating life, for traditional and activity based ventilation practice, the mine’s combined primary and auxiliary ventilation operating costs as well as the cost savings generated by an activity based ventilation system were determined and described in section 8 (for years 2008-2010), and Appendix B (for years 2011-2013, 2014-2017, 2018-2020). A summary with respect to the mine’s ventilation operating costs is provided in Table 26. Table 26: Mine A’s ventilation operating costs for traditional and activity based ventilation  Time Period  Traditional Ventilation ($/year)  Activity-Based Ventilation ($/year)  Savings in Operating Cost ($/year)  Savings in Operating Cost (%)  Years 2008-2010 Years 2011-2013 Years 2014-2017 Years 2018-2020  4,954,820 7,235,300 12,331,330 8,505,150  2,660,920 3,514,390 4,736,090 3,694,720  2,293,900 3,720,910 7,595,240 4,810,430  46 51 61 56  The overall economical and environmental benefits of an activity based versus traditional ventilation was determined and described in section 9. Besides the power consumption and associated operating costs, this analysis took into account the capital investments that are needed to develop and install an activity based ventilation control infrastructure at “Mine A”. 172  Simulations performed on Mine A’s ventilation model (years 2014-2017 configuration), showed that the intake air volume requirement during this time period will be beyond the capacity of the mine’s current ventilation infrastructure. Ventilation simulations also showed that to deliver this air volume through its primary system, the mine would need to replace eight main intake fans with larger fans and more powerful motors. Furthermore, preliminary calculations determined that the associated capital costs with respect to the replacement of 8 main intake fans would be uneconomical. An alternative solution proposed by the mine’s engineering department was to reduce the overall resistance of the mine’s primary system by constructing a new intake fresh air system. The capital investment of this new intake system was estimated at $3,154,680 of which $1,182,180 was allocated for year 2012 for the construction of the 4.2m diameter raisebore. The remaining of $1,972,500 would be used in 2013 for the purchase and installation of the Alphair surface fans and their heater house. With these capital investments accounted for, the potential benefits of the activity based ventilation controls system were determined for Mine A’s operating life. The economical analysis which compared the capital and operating costs of traditional versus activity based ventilation was performed for two ventilation scenarios, namely with and without the addition of the new intake system. The results were summarized in Table 21 and Table 22, respectively, which indicate the following: •  With traditional ventilation practice and an additional intake infrastructure, the combined capital and operating costs for the remaining life of “Mine A” would be $103.8M, on average $7.9M/year  •  With activity based ventilation practice, the combined capital and operating costs for the remaining life of “Mine A” would be $55.2M/year, on average $4.2M/year  •  The savings in capital and operating costs of an activity based ventilation control system during Mine A’s operating life would be $48.6M, which on average represents savings of $3.7M/year.  With traditional ventilation, as was expected, simulations showed that the new intake fresh air system would reduce the mine’s equivalent resistance. Furthermore, the additional new surface fans installed on top of the No. 5 shaft would also reduce the main intake fans’ operating pressure. However, these additional intake fans will have no influence on the main exhaust fans located on top of the No. 11 shaft, which would still operate at very high pressures. This tends to indicate that in addition to this new intake fresh air system, the mine 173  would need to construct an exhaust system as well. This would further increase the mine’s capital investments. However, if “Mine A” was to employ an activity based ventilation control system during its operating life, then no additional intake or exhaust infrastructure would be required. This concludes that at “Mine A”, the development and installation of an activity based ventilation control system would be far more effective and beneficial that any attempt to reduce the mine’s equivalent resistance through the construction of additional intake and exhaust infrastructure. For “Mine A”, the environmental benefits of an activity based ventilation control system were determined in section 9.3.2, and then summarized in Table 24. This table shows that during Mine A’s operating life, the combined saving in energy consumption would be in the range of 906.5 GWh, which represents average savings in energy consumption of 69.7 GWh/year. The savings in GHG emissions associated with the savings in energy consumptions were then expressed in tonnes of equivalent CO2. Calculations show that during Mine A’s operating life the savings in GHG emissions generated by an activity based ventilation control system would be in the range of 181,300 tonnes of equivalent CO2, which on average represents savings in GHG emissions of 13,947 tonnes/year of equivalent CO2.  174  11.  CONCLUSIONS  This thesis introduced a new evaluation method, where the magnitude of “ventilation redundancy” was used to measure the efficiency of two complex ventilation systems. This evaluation method was described in the case studies presented in section 5.3 and section 5.4, which explored the possibility of reducing the energy consumption in two large and deep underground metal mines. The case study presented in section 5.4 also showed that despite significant resources allocated for the determination of Kidd Creek mine’s intake air volume requirement, two conventional airflow evaluation methods did not agree. The first evaluation method, namely the “future production planning” analysis underestimated the mine’s intake air volume requirement because no additional airflows were allocated in order to prevent uncontrolled recirculation within the ventilation system and provide minimum airflow velocities throughout the inactive areas of the mine. The second evaluation method had the ability to capture the dynamic nature of the mine, however this method overestimated the mine’s intake air volume requirement due to the fact that the mining equipment did not have a “time stamp”. Consequently, fresh air was assigned to the same equipment entering the production level but at different times. To address these types of discrepancy in estimating airflow, this thesis introduces a novel method, where a multi-level mining block’s activity based intake airflow requirement was determined through discrete-event mining process simulations using AutoModTM. In accordance with the number of active mining blocks required to achieve future production requirements, the mine’s activity based life-cycle airflow demand schedule was subsequently determined. Furthermore, according to the mine’s activity based life-cycle demand schedule, its primary and auxiliary ventilation system was solved and balanced through ventilation simulation. Through post processing, using the fan activity data file generated by discrete-event process simulation performed on the “Mining Process” model, the “Fan Summary” macro determined that the multi-level mining block’s activity based intake air volume requirement is: Qactivityblock  = 73 m3/s. For the same mining block, but with traditional ventilation delivery, its intake  air volume requirement was calculated at: Qtarditional-block = 114.7 m3/s. This shows that an activity based ventilation control system could reduce the intake air volume requirement of 175  the mining block by 41.7 m3/s, which represents a 37% reduction. This reduction is comparable to reductions in intake airflow reported by other underground operations that already employ an activity-based ventilation control system. For example, at the Malmberget mine, operated by LKAB in northern Sweden, within the individual mining blocks that employ sublevel caving, the savings in intake airflow generated by an activity based ventilation system are reported to be in the range of 39% [Nensen & Ludkwist, 2005]. Furthermore, at the Creighton mine, operated by Vale Inco within the Sudbury basin, an activity based ventilation control system was installed on its 2340m production level. Average reductions in intake air volume generated by the auxiliary ventilation control system are reported to be in the range of 40%. Consequently, it can be stated that for model validation purposes, the above mentioned results show that the mining block’s activity based intake air volume requirement determined through discrete-event process simulation performed on the “Mining Process” model can be considered as being accurate. The economic and environmental benefits of this new ventilation design concept were analyzed on an existing large and deep metal mine, namely “Mine A”. During Mine A’s operating life, ventilation simulations showed that an activity based ventilation control system when compared to traditional ventilation, can reduce the intake air volume requirement of “Mine A” by 25% to 28%, which can be a function of the mine’s production output and mining depth. Consequently, the combined primary and auxiliary ventilation operating costs of “Mine A” could be reduced by 46% to 61 %. This shows that the reductions in Mine A’s intake air volume are less significant than the reductions in intake air volume within the individual mining blocks. This is due to the fact that at “Mine A” in addition to the production workings, fresh air needs to be delivered to the inactive areas of the mine in order to maintain minimum airflow velocities along their main levels and the ramp system. Calculations show that during the operating life of “Mine A”, an activity based ventilation control system could generate savings in energy consumption in the range of 906.5 GWh, which represents average energy savings of 69.7 GWh/year. To selectively deliver variable air volumes to the development and production workings based upon individual mining operations, it is essential to know where, when and for how long these air volumes are needed. In highly mechanized metal mines, this means knowing where production, development and service vehicles are located at any one time. This can be achieved through the use of a vehicle identification and tracking system. Such systems could 176  be designed to incorporate mining personnel identification and tracking as well. This could be achieved through the use of integrated communication cap lamps (ICLLs), which carry the Wi-Fi tags that can be identified and tracked through the local area networks (LANs) and the wireless access points (WAPs) of a mine’s communication system. This shows that besides reducing a mine’s energy consumption, an activity based ventilation control system and its communication infrastructure can also be used as an emergency notification and response system in case of underground fires and other emergency situations. In the past, ventilation system design was usually considered an “afterthought”, whereby a mine’s primary and auxiliary ventilation delivery systems were designed to fit predetermined mine layouts that were usually dictated by other factors such as a preselected mining method, extraction sequence, haulage systems or ground control considerations. This new ventilation design criterion could fundamentally change the way underground ventilation systems are presently designed an operated. This is because this new design criterion enables the integration of the ventilation system design and the overall mine design processes. For example, pre-feasibility and feasibility level projects such as Vale Inco’s Copper Cliff Deep project (CCD), or Xstrata Nickel’s Onaping Depth project, have recently used AutoModTM for production scheduling and mine planning purposes. This shows that the output data generated by discrete-event process simulation performed on mine models using AutoModTM for mine planning and design purposes are reliable. Consequently, the development of the “Mining Process” model and the determination of a mine’s activity based intake airflow requirement came as a natural extension of a design practice that has already been accepted by the mining industry.  177  12.  THESIS CONTRIBUTION AND FUTURE WORK  12.1 Thesis Contributions This thesis introduces a new method that can be used to evaluate the efficiency of large and complex underground ventilation systems. This new evaluation method is based upon the magnitude of a mine’s potential “ventilation redundancy” that can be used to gauge the efficiency of the mine’s ventilation system. The “ventilation redundancy” parameter was defined as the “total intake airflow delivered by the primary fans” minus the “sum of the activity based air volumes” delivered to the production workings and underground service facilities. This new evaluation method was tested and its efficacy demonstrated in two case studies: “Determination of Ventilation Redundancy at Creighton Mine - Vale Inco” and “Ventilation Utilization Efficiency at Kidd Creek Mine - Xstrata PLC”. This thesis introduces a novel method, which demonstrated that a mine’s activity based airflow requirement can be determined through discrete-event mining process simulation using AutoModTM. The “Fan Summary” macro described in section 6.2.4 has the capability to process the fan activity data generated by the discrete event process simulation into activity based air volume requirements. Furthermore, based upon the mine’s future production needs its life-cycle airflow demand schedule was then determined. With this new ventilation design concept, the accuracy of a mine’s “true” intake air volume requirement would only depend upon the depth of detail at which an underground operation is modeled and the accuracy to which the scheduled mining activities can be modelled using simulation. The use of “Mine A” as a case study demonstrated how the new method can be used to determine the capital, operating and environmental benefits of an activity based ventilation system designed according to the life-cycle airflow demand schedule of a mine, through ventilation simulations. The result of this compared well with results from existing activity based ventilation systems. Thus, the new method provides a mechanism for designers to evaluate the potential benefits of an activity based system versus a traditional system. This could allow more informed decisions to be made with respect to the type of the ventilation system and the most economical level of ventilation control that can be used at a particular mine. Finally using the approach developed in this thesis the ventilation design can accommodate various mine design elements such as a selected mining method(s), ore/waste haulage 178  systems or an extraction sequence which could be taken into account at the very initial stage of the mine planning and design process, thus eliminating unexpected future infrastructure upgrades.  12.2 Future Work In addition to the production workings, to incorporate the mine’s secondary service infrastructure and all stationary facilities (i.e. ore crushing, hoisting system, loading pockets, equipment refuelling stations), the development of a “mine-wide” model is recommended. The development of such a model would surely involve significant resources from both research organizations as well as from the mine’s engineering department. 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Circ. 7740, pp 13. 183  McPherson, M.J., 1993. “Subsurface Ventilation and Environmental Engineering”. Chapman & Hall, First Edition. McPherson, M.J., 1964. “Mine Ventilation Network Problems (solution by digital computer)” Colliery Guardian, August 21, pp. 253-254. Meets, E. J., Meyer, C.F., 1993. “Some Applications of Ductless Fans in South African Coal Mines”. Proceedings of the 6th U.S. Mine Ventilation Symposium, Salt Lake City, Utah, USA, June 21-23. Meyer, M.A., 2008. “Implementing a Tracking and Ventilation Control System at Barrick Goldstrike’s Underground Division”, Proceedings of the 12th US/North American Mine Ventilation Symposium (Ed: K.G. Wallace) ISBN: 978-0-615-20009-5, June 911, Reno, Nevada, USA. Mine Ventilation Services Inc., 1997. “ClimsimTM Simulator, Version 1.0”, Fresno, California, USA. Minutes of the Meeting of the Canadian ad-hoc Diesel Committee, 1990. Sudbury, Ontario, Canada, December. National Institute of Occupational Health and Safety, 1988. “Carcinogenic Effects of Exposure to Diesel Exhaust”. Current Intelligence Bull. 50, Department of Health and Human Services (NIOSH). Publication 88-116, 30pp. Natural Resources Canada, 1998. “Canada’s Energy Outlook for 1996 – 2020”, http://nrcan.gc.ca/es/ceo/toc-96E.html. Natural Resources Canada, 2002. “List of CANMET-MMSL Approved Diesel Engines in Accordance with CSA Standards”, M424.2-M90 & M424.1-88. Nensen, P., and Lundkvist, A., 2005. “From 167 GWh to 72 GWh – Ventilation on Demand in LKAB’s Iron Ore Mine, Malmberget”, Proceedings of the IPPC Conference, Innovative Energy Efficiency Examples of Different Industrial Sectors – Energy Efficiency in Cement, Metal and Petrochemical Industry, Vienna, Austria, http://www.umweltbundesamt.at/fileadmin/site/uweltthemen/industrie/IPPC_Konfere nz/Nensen.pdf. O’Connor, D.F., 2008. “Ventilation on demand (VOD) auxiliary fan project – Vale Inco Limited, Creighton Mine”, Proceedings of the 12th US/North American Mine Ventilation Symposium (Ed: K.G. Wallace), pp. 41-44, June 9-11, Reno, Nevada, USA, ISBN: 978-0-615-20009-5 Partyka, J, Hardcastle, S.G., Kocsis, C., 1998. “Computer Simulation of Climatic Conditions for Automated Underground Metal Mines for Standard and Reduced Airflow Requirements”. Proceedings on CD Published by University of MissouriRolla, Rolla, Missouri, USA.  184  Payne, T., Mitra, R., 2009. “A Review of Heat Issues in Underground Metalliferous Mines”, Proceedings of the 12th US/North American Mine Ventilation Symposium, (Ed: K.G. Wallace), June 9-11, Reno, Nevada, pp. 197-201, ISBN: 978-0-615-20009-5. Rohrer M.W., McGregor I.W., 2002. “Simulating Reality Using AutoModTM”. Proceedings of the 2002 Winter Simulation Conference, December 8 – 11, San Diego, California Sarin, N., Gangal, M., & Feres, V., 1997. “Diesel Engine Performance for Reduced Ventilation in a Simulated Mine Environment”. CANMET-Mining and Mineral Sciences Laboratories, Natural Resources Canada, Divisional Report. Schutte, P.C., Marx, W.M, 2007. “Heat Stress Management in Deep and Hot Mines”, Published by Australian Centre for Geomechanics (ACG), Editors: Y. Potvin, J. Hadjigeorgiou, T.R. Stacey, Perth, Western Australia, ISBN 978-0-9804185-1-4, pp. 551-554. Schutte, P.C., Kielblock, A.J., 1997. “Heat Stress Protection in Abnormally Hot Environments”, Proceedings of the 6th International Mine Ventilation Congress, (Editor: R.V. Ramani), Pittsburgh, Pennsylvania, pp. 285-189, May 17-22. Scott, D.R. and Hinsley, F.B. (1951-1952). “Ventilation Network Theory, Parts 1 to 5”. Colliery Eng. 18, 29. Stachulak, J.S., 1989. “Ventilation Strategy and Unique Air Conditioning at Inco Limited”. Proceedings of the 4th U.S. Mine Ventilation Symposium, Berkeley, California, (Ed. M.J. McPherson), SME, Littleton, Col., ISBN 0-87335-082-0, pp. 3-9, June 5-7. Tonnos, A.M., Allen, C., 2008. “Technology Convergence for Sustainable Underground Mine Ventilation System Control”, Proceedings of the 12th US/North American Mine Ventilation Symposium (Ed: K.G. Wallace), pp. 37-40, June 9-11, Reno, Nevada, USA, ISBN: 978-0-615-20009-5. U.S. Environmental Protection Agency, 2000. “Health Assessment Document for Diesel Exhaust”. Report EPA/600/8-90/057E, July. Vale Inco, 2004. “Private Communication with Creighton Mine’s Engineering Department Vale Inco” Vagenas Nick, 1999. “Applications of Discrete-Event Simulation in Canadian Mining Operations in the Nineties”. International Journal of Surface Mining, Reclamation and Environment, Balkema, 3000 BR Rotterdam, Netherlands, pp 77-78. Vagenas, N., Scoble, M, Corkal, T. and Baiden, G., 1998. “Simulation of Vertical Retreat Mining Using AutoMod”. 27th International Symposium on Applications of Computers and Operations research in the Mining Industry (APCOM), April 19-23, London, England, ISBN 1870706366. Wallace, K., 2002. “General Operational Characteristics and Industry Practices of Mine Ventilation Systems”. Proceedings of the 7th International Mine Ventilation Congress, 185  Krakow, Poland, (Ed. S. Wasilewski), pp 229-236, (EMAG, Poland), ISBN 83913109-1-4. Wallace, K.G., Tessier, M., Pahkala, M., and Sletmoen, L., 2005. “Ventilation Planning at the Red Lake Mine”, Proceedings of the 8th International Mine Ventilation Congress (Ed: Gillies), pp 447-455, Australian Institute of Mining and Metallurgy, Melbourne. Wallace K.G., 2001. “General Operational Characteristics and Industry Practices of Mine Ventilation Systems”, Proceedings of the 7th International Mine Ventilation Congress (Ed: S. Wasilewski), pp 229-234, (Research and Development Centre for Electrical Engineering and Automation in Mining, Cracow, Poland.  186  APPENDICES Appendix A A.1  Modelling and Analysis Work Performed by the Candidate versus Contribution and Assistance from Others  A.1.1 Case Study #1: Determination of Ventilation Redundancy at Creighton Mine -Vale Inco Work performed by the candidate: •  Throughout a 3-week time period at the mine site, collected ventilation data to update Creighton Mine’s primary ventilation system. Explored the underground production levels and examined production and support operations to understand the primary and auxiliary airflow delivery practice at the mine as well as its operating characteristics (shift duration, haulage system, mining method, stope sequence, ground control, mining equipment, blasting practice, other support services).  •  Developed Creighton mine’s primary ventilation system from surface to 7400 level (2,250m depth) using VnetPCTM software.  •  Solved and balanced the ventilation model. Validated the ventilation model by comparing ventilation data generated trough ventilation simulation (i.e. airflow distribution, fan operating duties, pressure losses) versus airflow and differential pressures values that were measured underground. This validated ventilation model became the mine’s baseline model for further analysis.  •  Based upon future mining plans provided by the mine’s engineering department, extended the ventilation model of Creighton mine from 7400 level to the 8200 level (2,500m depth). This would represent the future configuration of the mine’s primary ventilation infrastructure for year 2019.  •  During the production activity monitoring time frame (125 days), installed temperature and relative humidity data loggers within two production draw-point areas as well as backfilling, drilling and ore-pass sites. A sample showing the locations where these data loggers were installed around a production stope area is shown in Figure 6, section 5.3.2.  •  During the monitoring time period downloaded temperature and relative humidity data onto a mobile computer and processed these data into dry-bulb temperature and 187  relative humidity graphs. Two samples of these graphs are presented in section 5.3.2 of the thesis (Figure 7 and Figure 8). •  Through ventilation simulation determined the mine’s intake air volume requirement for two levels of ventilation control, namely “time-of-day” and “activity-based”. This led to the determination of the mine’s potential ventilation redundancy within its primary and auxiliary ventilation systems based upon these two levels of control.  •  Based upon the mines potential intake air volume reduction, performed a detailed economic analysis and through ventilation simulation determined Creighton mine’s savings in installed fan power (kW), energy consumption (kWh/year) and operating cost ($/year) for two level of airflow control, “time-of-day” and “activity based”. This analysis was performed for two time periods, years 2003 – 2007 and 2015 – 2019. Table 6, Table 7 & Table 8 presented in section 5.3.5 were developed based upon data carried out through this economic analysis.  Work performed by others: Douglas O’Connor - Ventilation Supervisor at Creighton Mine •  Assisted the candidate in validating the baseline model of Creighton mine’s primary ventilation system and provided transportation to download the climatic data from the monitoring units (e.g. dry-bulb temperature, relative humidity, barometric pressure) and for ventilation measurements. The ventilation supervisor from Creighton mine provided the candidate historical ventilation data (e.g. airflow, pressure profiles) which were measured at the mine over the course of several years. For validation purposes the candidate compared these data with measurements performed by himself and the output data generated by the mine’s ventilation model obtained through simulation.  Stephen Hardcastle - CANMET-MMSL •  Analyzed the production blast monitoring at Creighton mine and the decaying profiles of CO, CO2, NO and NO2 gases. Determined the longhole blasting fume clearance times based upon CO decaying profile.  •  Based upon the individual temperature and relative humidity graphs carried out by the candidate, constructed the long-term mucking, drilling and backfilling logs. Based  188  upon these long-term activity logs, identified the time frame during which the mine’s primary ventilation system can operate at a reduced level. •  Reviewed the ventilation modelling work and the economic analysis performed by the candidate.  A.1.2 Case Study #2: Ventilation Utilization Efficiency at Kidd Creek Mine – Xstrata PLC Work performed by the candidate: •  During a 2-week time period at the Kidd Creek complex (Timmins, Ontario) went on two underground visits to explore and understand the mine’s primary and auxiliary ventilation delivery system as well as its operating characteristics (e.g. mining cycle, mining method, stope sequence, ground control, blasting practice, type and size of production & service equipment).  •  Based upon ventilation data provided by the engineering department, developed the ventilation model of Kidd Creek mine, which includes “No. 1 Mine”, “No. 2 Mine”, “No. 3 Mine” and “Mine D” from surface through to Level 88 located at 3,000m depth using VnetPCTM software. Solved and balanced the mine’s primary ventilation system.  •  Developed a macro that was able to extract and transfer production, development, backfilling and miscellaneous activity as well as all associated production equipment data from the mine’s SIMS database into independent MSExcelTM spreadsheet files.  •  Within MSExcelTM and through “pivot table” type analysis determined the mine’s active production and development levels for various time periods such as for one month, one week and for one day.  •  Through the same “pivot table” type analysis identified the production equipment that were operating within the active production and development levels (e.g. LHDs, haulage trucks, drills) and based upon the size of the equipment determined the air volume requirements for each active level of the mine.  •  Assisted in evaluating Kidd Creek mine’s intake airflow requirement as a function of its airflow redistribution frequency on the active levels (monthly, weekly and daily) and consequently the determination of its potential ventilation redundancy. Samples of these analysis are shown in Table 10, Table 11 (section 5.4.3) 189  Work performed by others: Gary Li - CANMET-MMSL •  Assisted the candidate with the “pivot table” analysis method, using MSExcelTM.  Stephen Hardcastle - CANMET-MMSL •  Reviewed the work performed by Genivar Engineering which was based upon the mine’s 18-month production plan (second stage – iteration #1).  •  Modified the air volumes determined by Genivar Engineering for production, drilling and miscellaneous activities in order to eliminate potential airflow recirculation within the mine’s auxiliary ventilation system (second stage – iteration #2)  •  Based upon the pivot table analysis performed by the candidate, calculated the mine’s intake airflow requirement as a function of its airflow redistribution frequency. Sample of this calculation is shown in Figure 11 (section 5.4.3)  A.1.3 Mining Process Model Development Using AutoModTM Simulator Work performed by the candidate: •  Collected mining activity data (time occurrence/duration) from an underground operation that employs the vertical retreat mining (VRM) method. With the mining activity data entered in the “Mining Process” model through its “Data Input” utility, performed discrete-even process simulation trials in order to determine the activity based intake air volume requirement of the mining block. Early simulation trials showed that the mining operations within various production cycles did not match the operation sequence observed in an existing mine that employs the VRM method. In addition, the model initiation functions of the “Mining Process” model did not have the capability to take into account scheduled (e.g. lunch breaks) and unscheduled (e.g. equipment failure) production down-times.  •  Eliminated all global variables in the “Mining Process” model and defined new local variables for the mining operations associated with the production stopes, developments and ore/waste passes. Modified the process logic of the simulation model to accommodate the newly defined local variables.  •  Within the process system of the “Mining Process” model, developed new activity subroutines for drilling, explosive loading, blasting and ore/waste transportation to 190  synchronize the sequence of mining operations simulated by the “Mining Process” model with the operating sequence observed at an existing mine that employs VRM. •  Within the process system, developed a new process subroutine saved in an independent source file that has the ability to stop and re-initiate the mining operations that are affected by scheduled or unscheduled production down-times. The local ventilation variables that were defined in this new process subroutine have the capability of taking into account scheduled and unscheduled production down-times such as lunch breaks, shift changes, blasting fume clearances, weekend shutdowns and equipment failures when determining the mining block’s activity based intake air volume requirement.  •  Developed a new macro that has the ability to transfer mining activity (time occurrence/duration), equipment data (type/size), stope capacity as well as production down-times entered by the user in the “Data Input” utility to the model initiation functions of the “Mining Process” model. Defined new model initiation functions to transfer these data to the various subroutines of the process logic.  •  Developed the “Fan Volume” worksheet through which the operating duties of the auxiliary fans assigned to the ventilation system of the “Mining Process” model are entered and transferred to the “Fan Summary” macro.  •  Redesigned the “Fan Summary” macro to enhance its functionality and capability of taking into account any additional air volumes that could be required to maintain minimum airflow velocities along the access ramp and potential inactive areas of the mining block. For the fans that supply fresh air during operations that immediately follow blasting fume clearance time, the “Fan Summary” macro has the capability to take into account “early fan start” times as well.  •  Trough discrete event process simulation performed on the “Mining Process” model determined the mining block’s “activity-based” air volume requirement versus “traditional” ventilation demand.  Work performed by others: Penguin Automated Systems Inc. •  Developed the initial version of the “Mining Process" model using AutoModTM, which includes one “process system” and 5 “path moving” systems. The path moving 191  systems graphically represent a typical multi-level mining block that employs the VRM method consisting of two production stopes, two developments, their associated ore & waste passes and a ramp system. Penguin Automated Systems Inc was hired to provide technical support only, namely for coding using AutoModTM language. The concept and structure of the “Mining Process” model, its sub-components and the interaction logic between the sub-components and the mining block’s ventilation system were deigned by the candidate and incorporated in the simulation model under the guidance of the candidate. Based upon discrete-event process simulation trials, the candidate then modified the process system of the model, designed and developed new process subroutines in order to match the sequence of the mining operations simulated by the “Mining Process” model with the operating sequence observed at an existing mine. Gary Li – CANMET-MMSL •  Assisted in the development of the “Fan Summary” macro, which was used to manipulate and post-process fan activity output data generated by discrete-event mining process simulations performed on the “Mining Process” model.  A.1.4 Underground Operations Survey to Identify the Canadian Mines that Would Benefit the Most from an Activity Based Ventilation Control System Work performed by the candidate: •  Performed mining activity reviews for underground operations within Canada. The activity reviews focused on the employed mining method, haulage systems, mining sequence and ground control practices. Visited various underground operations in the province of Ontario and Quebec.  •  Classified the Canadian mining operations based upon the most frequently employed extraction method.  •  Identified the underground operations that would benefit the most from the new ventilation design criteria.  Work performed by others: • No input  192  A.1.5 Determination of Short-Term and Life-Cycle Airflow Demand Schedule Work performed by the candidate: •  For practical purposes and to accurately determine to potential benefits of this new design concept, the intake air volume was determined for an existing multi-level mine, namely “Mine A” which employs the VRM method.  •  Based upon current ventilation design criteria, determined Mine A’s short-term intake air volume requirement. According to the number of active stopes that are needed to achieve production targets, determined the life-cycle airflow demand schedule of “Mine A” for “traditional” ventilation.  •  Based upon the output data generated through discrete-event process simulation performed on the “Mining Process” model and according to the number of active stopes that are needed to achieve production targets, determined the “life-cycle” airflow demand schedule of “Mine A” for “activity-based” ventilation.  Work performed by others: • No input  A.1.6 Ventilation System Design Based Upon the Life-Cycle Airflow Demand Schedule Work performed by the candidate: •  Based upon ventilation data and future mining plans provided by its engineering department, developed the model of Mine A’s primary ventilation system for four time periods, namely, years 2008-2010, 2011-2013, 2014-2017 & 2018-2020, using VnetPCTM software.  •  For each time period mentioned above, solved and balanced the primary ventilation system of “Mine A” for “traditional” and “activity-based” ventilation.  •  For each time-period, through ventilation simulation, determined the fan power (kW), energy consumption (MWh/year) and operating cost ($/year) of Mine A’s primary ventilation system. This was performed for “traditional” and “activity-based” ventilation requirements.  •  Through ventilation simulation, determined the fan power (kW), energy consumption (MWh/year) and operating cost ($/year) of Mine A’s auxiliary ventilation system.  193  Again, this was performed for “traditional” and “activity-based” ventilation requirements. •  For “Mine A” and for each time period mentioned above, determined the savings in energy consumption (MWh/year) and operating costs ($/year) of an “activity-based” ventilation system versus “traditional” ventilation. Table 16 (section 8.3) was developed based upon the data generated through ventilation simulation.  Work performed by others: • No input  A.1.7 Economic and Environmental Benefits of this New Ventilation Design Approach Work performed by the candidate: •  Performed budgetary cost estimate of an “activity-based” ventilation control system at “Mine A” (private communication with Bestech Engineering)  •  Performed a cost estimate for a new intake fresh air infrastructure at “Mine A”, to lower the operating pressure of the mine’s the primary fans (private communication with HATCH Engineering)  •  Determined the economic and environmental benefits of an “activity-based” ventilation system at “Mine A” versus “traditional” ventilation practice (operating costs, energy consumption, GHG emissions in equivalent CO2).  Work performed by others: • No input  194  Appendix B B.1  Solving and Balancing the Primary and Auxiliary Ventilation Systems of “Mine A” for Years 2011–2013, Years 2014-2017, Years 2018-2020 with Traditional Air Volume Requirements  B.1.1 Ventilation Modelling for Years 2011–2013 (Traditional Ventilation) Long-term planning data show that for years 2011 – 3013, the production requirements at “Mine A” will increase from 4,000 tons/day to 5,000 tons/day. As the capacity of a mining block is 1,000 tons/day, to fulfill production requirements, the mine would need to maintain 5 mining blocks active. Within this time period the production stopes will be located on the following levels: •  Mining block No. 1 with stopes on the 7400 and 7530 levels  •  Mining block No. 2 with stopes on the 7660 and 7720 levels  •  Mining block No. 3 with stopes on the 7780 and 7840 levels  •  Mining block No. 4 with stopes on the 7900 and 7960 levels, and  •  Mining block No. 5 with stopes on the 8020 and 8080 levels  Primary System – Flow Distribution, Fan Duties and Fan Power Requirements The primary ventilation model for this time period was developed by updating the previous ventilation model (for years 2008 – 2010) with additional levels as the depth of the development and production workings increase. According to the mine’s long-term planning data, the additional production levels that were added to the previous model were the 7900, 7960, 8020 and 8080 levels. Between years 2011 – 2013, the mines total intake air volume (as determined in section 7.2.1), which includes the air volumes required by production and secondary services is: QInatake(2011-2013) = Qproduction-traditional + Qservices = 575 m3/s + 200 m3/s = 775 m3/s. As previously mentioned, the total air volume for secondary services (Qservices) includes the air volumes required by backfilling and ore crushing activities and the air volumes delivered to the inactive areas of the mine. Based on these flow requirements the mine’s primary ventilation system was solved and balanced through ventilation simulations. The airflow distribution within the primary ventilation system is shown in Figure 45. 195  Ventilation simulations also show that of its total intake air volume of 775 m3/s, approximately 170 m3/s is delivered down the No. 9 intake shaft. This air is then distributed along the inactive levels located within the upper orebody of the mine, namely between the 3800 level and 7000 level. The fresh air needed by the production workings located between the 7200 level and 8080 level, is delivered down the main FAR and the No. 8 internal shaft, namely 350m3/s and 234m3/s, respectively. Approximately 759 m3/s of contaminated air returning from the upper levels and the production workings is directed to surface through the No. 11 exhaust shaft. Another 16 m3/s of contaminated air resulting from secondary activities performed on the 3800 level is directed to surface through the No. 5 exhaust shaft (see Figure 45). Within the primary system, the operating characteristics of the main fans in terms of air quantity (Q) and total pressure (Ptotal) are: •  1420 Level Intake Fans (2 in parallel): Q = 170 m3/s, Ptotal = 1,390 Pa  •  2300 Level Intake Fans (2 in parallel): Q = 350 m3/s, Ptotal = 3,285 Pa  •  2600 Level Intake Fans (2 in parallel): Q = 234 m3/s, Ptotal = 3,008 Pa  •  No. 11 Shaft Exhaust Fans (3 in parallel): Q = 759 m3/s, Ptotal = 5,500 Pa  The operating characteristics and the fan power of all individual primary fans (main and boosters) that operate within the primary system are shown in Table 27. Between years 2011 – 2013, the total fan power within the primary ventilation system is: PPrimary = 7,473 kW. For average 70% fan efficiency and for $0.055/kWh electricity charge, the primary ventilation system operating cost is: OCPrimary(2011-2013) = 7,473 kW/0.7 x 365 days/year x 24 hours/day x $0.055/kWh = $5,143,600/year.  196  Figure 45: Airflow distribution and fan operating duties in the primary system for years 2011 – 2013 - Traditional ventilation  197  Table 27: Fan operating duties and power requirements for primary system (years 2011 - 2013) – Traditional ventilation Junction From 191 383 39 380 420 487 7722 7726 7788 7784 7853 7845 7851 7908 7969 7905 8032 8090 7965 8025 8083 13 53 232 9 6 TOTAL  Junction To 190 384 40 381 421 485 7721 7727 7789 7783 7854 7844 7852 7909 7970 7906 8033 8091 7966 8026 8084 29 54 58 10 5  Pressure (kPa) 0.088 0.825 1.687 0.571 2.293 0.111 0.029 0.172 0.186 0.038 0.364 0.113 0.222 0.403 0.420 0.043 0.429 0.404 0.062 0.075 0.068 1.900 1.390 3.285 3.008 5.500  Flow Fan Power Operating Cost Fan Curve (m3/s) (kW) ($/year) Status 5.9 0.52 358 On 10.1 8.29 5,705 On 350.0 590.41 406,374 On 20.5 11.71 8,061 On 3.2 7.27 5,007 On 18.5 2.05 1,412 On 23.1 0.67 461 On 12.7 2.18 1,499 On 12.3 2.28 1,568 On 27.9 1.06 729 On 73.2 26.66 18,349 On 72.7 8.22 5,657 On 8.9 1.98 1,365 On 71.9 28.97 19,943 On 73.8 31.00 21,335 On 10.1 0.43 299 On 73.9 31.71 21,828 On 11.8 4.78 3,290 On 8.6 0.53 367 On 11.2 0.84 577 On 32.4 2.20 1,515 On 234.0 444.59 306,007 On 170.0 236.27 162,624 On 350.0 1,149.74 791,352 On 234.0 703.86 484,457 On 759.0 4,174.42 2,873,196 On 7,473.00 5,143,600  Fan Arrangement 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 2 in Parallel 2 in Parallel 2 in Parallel 3 in Parallel  Fan Location 1900 LEVEL 5 SHAFT 3800 LEVEL 5 SHAFT 5400L Fans 3800L Booster Fan 4200L Booster Fan 4800L Booster Fan 7660 FAR booster fan 7720L Booster fan 7780L Booster Fan 7660 FAR Booster 7840L Booster fan 7660 FAR Booster 7840 Esc. Booster 7900L Booster fan 7960L Booster Fan 7900L booster 8020L Booster fan 8080L Booster Fan 7960L booster 8020L booster 8080L booster 5000L Fans (2p) 1420L Fans (2p) 2300L Fans (2p) 2600L Fans (2p) 11 Shaft Fans (3p)  198  Auxiliary System – Flow Distribution, Fan Duties and Power Requirements At “Mine A”, the auxiliary ventilation setup within each mining block is fairly similar; the fresh air to the production and development faces is delivered through steel ducts under the assistance of auxiliary fans with motor powers that range between 25 to 75 kW. Furthermore, the size of diesel equipment used for production, development and haulage operations is similar. As a result, the air volume requirements within each mining block remain the same. Throughout this time-phase analysis, the fan operating duties and the airflow distribution within each mining block is similar (see Figure 35). A detailed description of the auxiliary ventilation parameters was already provided in section 8.1.1. With respect to the auxiliary ventilation system, during this time-phase analysis, the total air volume required by all production activities will depend upon the number of mining blocks that would need to be maintained active in order to fulfill production demand. For traditional ventilation practice, the duty and the operating power of each auxiliary fan assigned to one mining block was already shown in Table 13. This table also showed that the total fan power within one mining block was: PBlock = 299 kW. Between years 2011 - 2013 the long-term air volume schedule presented in section 7.2.1 shows that to fulfill production requirements the mine would need to maintain active 5 mining blocks. As a result, the total fan power for the production activities is: PProduction = 299 kW/mining block x 5 mining blocks = 1,495 kW. For secondary services and to provide minimum air volumes throughout the inactive areas the mine uses 62 additional auxiliary fans. The total fan power for the secondary services is: PServices = 24.9 kW x 62 fans = 1,544 kW. For this time period the combined fan power of the mine’s auxiliary system is: PAuxilairy = PProduction + PServices = 1,495 kW + 1,544 kW = 3,039 kW. For average 70% fan efficiency and for $0.055/kWh electricity charge, the secondary ventilation system operating cost is: OCSecondary = 3,039 kW/0.70 x 365 days/year x 24 hours/day x $0.055/kWh = $2,091,700/year. Between years 2011 – 2013 and with traditional ventilation practice, the mine’s total fan power is: PTotal(2011-2013) = PPrimary = PAuxilairy = 7,473 kW + 3,039 kW = 10,512 kW. For an average 70% primary and auxiliary fan efficiency and for $0.055/kWh electricity charge, the total ventilation operating cots at “Mine A” is: OCTotal(2011-2013) = OCPrimary(2011-2013) + OCSecondary(2011-2013) = $5,143,600/year + $2,091,700/year = $7,235,300/year. 199  B.1.2 Ventilation Modelling for Years 2014–2017 (Traditional Ventilation) Long-term planning data shows that between years 2014 – 2017 the production requirements at “Mine A” will further increase from 5,000 tons/day to 6,000 tons/day. As the capacity of one mining block is 1,000 tons/day, to fulfill production requirements the mine would need to maintain 6 mining blocks (12 stopes) active. To achieve this, the mine would need to access and develop two additional production levels, namely the 8140 (2,482m) and 8200 (2,500m) levels. Each mining block includes two stopes located on the following levels: •  Mining block No. 1 with stopes on the 7400 and 7530 levels  •  Mining block No. 2 with stopes on the 7660 and 7720 levels  •  Mining block No. 3 with stopes on the 7780 and 7840 levels  •  Mining block No. 4 with stopes on the 7900 and 7960 levels  •  Mining block No. 5 with stopes on the 8020 and 8080 levels, and  •  Mining block No. 6 with stopes on the 8140 and 8200 levels  Primary System – Flow Distribution, Fan Duties and Fan Power Requirements The primary ventilation model for this time period was developed by updating the ventilation model for years 2011 – 2013 with additional levels as development and production workings deepen. According to long-term mine planning data, the additional production levels that were added to the previous model were the 8140 (2,482m) and the 8200 (2,482m) levels. Between years 2014 – 2017 the mine’s total intake air volume (as determined in section 7.2.1), which includes the air volumes required by production and secondary services is: QInatake(2014-2017) = QProduction-traditional + QServices = 690 m3/s + 200 m3/s = 890 m3/s. Based on these flow requirements the mine’s primary ventilation system was solved and balanced by means of ventilation simulation. The airflow distribution within the primary ventilation system is shown in Figure 46. Ventilation simulations show that these air volume requirements are well beyond the capacity of the mine’s primary ventilation system. With the mine’s current intake infrastructure, to deliver these air volumes throughout its primary system, the main surface fans (especially the fans located on top of the No. 11 exhaust shaft) would need to operate at high pressures. To reduce the mine’s resistance and consequently the fan pressure requirements, the mine would need to construct an additional intake system. The economical impact of this additional intake system is discussed in section 9.  200  Figure 46: Airflow distribution and fan operating duties in the primary system for years 2014 – 2017 - Traditional ventilation  201  The main fans operating duties of the primary ventilation system are: •  1420 Level Intake Fans (2 in parallel): Q = 145m3/s, Ptotal = 3,800 Pa  •  2300 Level Intake Fans (2 in parallel): Q = 438m3/s, Ptotal = 4,576 Pa  •  2600 Level Intake Fans (2 in parallel): Q = 293m3/s, Ptotal = 4,394 Pa  •  No. 11 Shaft Exhaust Fans (3 in parallel): Q = 874m3/s, Ptotal = 8,325 Pa  The operating duties and the fan power of all individual primary fans (main and boosters) that operate within the primary system are shown in Table 28. Between years 2014 – 2017, the total fan power within the primary ventilation system is: PPrimary = 14,578 kW. For an average 70% fans efficiency and for $0.055/kWh electricity charge, the primary ventilation system operating cost is: OCPrimary (2014-2017) = 14,578 kW/0.7 x 365 days/year x 24 hours/day x $0.055/kWh = $10,033,830/year. Auxiliary System – Flow Distribution, Fan Duties and Power Requirements For years 2014 – 2017 time-period the auxiliary ventilation setup is similar to the previous scenarios. The operating duty of the auxiliary fans and the airflow distribution within one mining block were presented in Figure 35. Based upon traditional ventilation practice, the combined fresh air required at one mining block is QTotal-block = 114.7m3/s. Table 13 shows that the combined fan power within one mining block is PBlock = 299 kW. The long-term air volume schedule from section 7.2.1 shows that to fulfill production requirements the mine would need to maintain 6 mining blocks active. Consequently, the total fan power for production activities is: PProduction = 299 kW/mining block x 6 mining blocks = 1,794 kW. The total fan power for secondary services would remain the same at: PServices = 24.9 kW x 62 fans = 1,544 kW. The combined fan power of the mine’s auxiliary system is: PAuxiliary = PProduction + PServices = 1,794 kW + 1,544 kW = 3,338 kW. For an average 70% fan efficiency and for $0.055/kWh electricity charge, the secondary ventilation system operating cost is: OCSecondary = 3,338 kW/0.70 x 365 days/year x 24 hours/day x $0.055/kWh = $2,297,500/year.  202  Table 28: Fan Operating Duties and Power Requirements for Primary System (Years 2014 - 2017) – Traditional ventilation Junction Junction From To 7726 7727 7788 7789 7853 7854 7908 7909 7969 7970 8090 8091 8152 8153 8210 8211 8032 8033 8205 8206 8208 8209 383 384 13 29 39 40 191 190 487 485 36 35 7844 7843 7851 7852 7906 7907 7965 7966 8025 8026 8146 8147 8083 8084 53 54 232 58 9 10 6 5 TOTAL  Pressure (kPa) 0.669 0.684 0.311 0.313 0.346 0.413 0.452 0.544 0.365 0.260 0.270 2.500 4.200 4.400 0.100 0.127 1.973 0.431 0.217 0.044 0.085 0.121 0.198 0.160 3.800 4.576 4.394 8.325  Flow Fan Power Operating Cost Fan Curve Fan (m3/s) (kW) ($/year) Status Arrangement 16.21 10.84 7,464 On 1 in Series 25.56 17.48 12,033 On 1 in Series 106.23 33.04 22,738 On 1 in Series 46.80 14.65 10,083 On 1 in Series 1.73 0.60 413 On 1 in Series 63.11 26.06 17,939 On 1 in Series 50.85 22.98 15,819 On 1 in Series 56.83 30.91 21,277 On 1 in Series 15.06 5.50 3,783 On 1 in Series 27.63 7.18 4,945 On 1 in Series 15.41 4.16 2,863 On 1 in Series 8.82 22.06 15,185 On 1 in Series 293.00 1,230.58 846,994 On 2 in Parallel 438.01 1,927.24 1,326,489 On 1 in Series 4.51 0.45 311 On 1 in Series 19.22 2.44 1,680 On 1 in Series 34.98 69.01 47,501 On 1 in Series 30.45 13.12 9,034 On 1 in Series 26.57 5.77 3,969 On 1 in Series 96.76 4.26 2,930 On 1 in Series 47.20 4.01 2,762 On 1 in Series 19.83 2.40 1,652 On 1 in Series 10.89 2.16 1,484 On 1 in Series 11.82 1.89 1,302 On 1 in Series 145.00 551.00 379,246 On 2 in Parallel 438.00 2,004.31 1,379,537 On 2 in Parallel 293.00 1,287.43 886,117 On 2 in Parallel 874.00 7,276.06 5,008,006 On 3 in Parallel 14,578.00 10,033,830  Fan Location 7720L Booster Fan 7780L Booster Fan 7840L Booster Fan 7900L Booster Fan 7960L Booster Fan 8080L Booster Fan 8140L Booster Fan 8200L Booster Fan 8020L Booster Fan 8200L booster 8200L booster fan 3800L 5 Shaft fan 5000L Fans (2p) 5400L Fans 1900L 5 Shaft fan 6 shaft xc on 4800 L FAR 7200-7400 7840L fan 7840L to FAR/Esc. 7900L fan 7960L fan 8020L fan 8140L to ramp 8080L fan 1420L Fans (2p) 2300L Fans (2p) 2600L Fans (2p) 11 Shaft fans (3p)  203  For years 2014 – 2017 and with traditional ventilation practice, the mine’s total fan power is: PTotal(2014-2017) = PPrimary = PAuxiliary = 14,578 kW + 3,338 kW = 17,916 kW. For average 70% fan efficiency and for $0.055/kWh electricity charge, the total ventilation operating cots at “Mine A” is: OCTotal(2014-2017) = OCPrimary(2014-2017) + OCSecondary(2014-2017) = $10,033,830/year + $2,297,500/year = $12,331,330/year.  B.1.3 Ventilation Modelling for Years 2018–2020 (Traditional Ventilation) Between years 2018 – 2020 long-term mine planning data shows that production requirements at “Mine A” will decrease from 6,000 tons/day to 5,000 tons/day. With the capacity of 1,000 tons/day per mining block, the mine would need to maintain 5 mining blocks active in order to fulfill production demand. Each mining block has two stopes located on the following levels: •  Mining block No. 1 with stopes on the 7530 and 7660 levels  •  Mining block No. 2 with stopes on the 7720 and 7780 levels  •  Mining block No. 3 with stopes on the 7900 and 7960 levels  •  Mining block No. 4 with stopes on the 8020 and 8080 levels, and  •  Mining block No. 5 with stopes on the 8140 and 8200 levels  Primary System – Flow Distribution, Fan Duties and Fan Power Requirements Between years 2018 – 2020 the mine’s total intake air volume (as determined in section 7.2.1), which includes the air volumes required by production and secondary services is: QInatake(2018-2020) = QProduction-traditional + QServices = 575 m3/s + 200 m3/s = 775 m3/s. The airflow distribution within the primary ventilation system is shown in Figure 47. The main fans operating duties of the primary ventilation system are: •  1420 Level Intake Fans (2 in parallel): Q = 143m3/s, Ptotal = 2,944 Pa  •  2300 Level Intake Fans (2 in parallel): Q = 369m3/s, Ptotal = 3,807 Pa  •  2600 Level Intake Fans (2 in parallel): Q = 248m3/s, Ptotal = 3,587 Pa  •  No. 11 Shaft Exhaust Fans (3 in parallel): Q = 760m3/s, Ptotal = 6,342 Pa  The operating characteristics and the operating power of all individual fans (main and boosters) that are part of the mine’s primary ventilation system are shown in Table 29.  204  Figure 47: Airflow distribution and fan operating duties in the primary system for years 2018 – 2020 - Traditional ventilation  205  Table 29: Fan Operating Duties and Power Requirements for Primary System (Years 2018 - 2020) – Traditional ventilation Junction From 7726 7788 7853 7908 7969 8090 8152 8210 8032 8205 8208 383 13 39 191 487 36 7844 7851 7906 7965 8025 8146 8083 232 53 9 6 TOTAL  Junction To 7727 7789 7854 7909 7970 8091 8153 8211 8033 8206 8209 384 29 40 190 485 35 7843 7852 7907 7966 8026 8147 8084 58 54 10 5  Pressure (kPa) 0.669 0.684 0.311 0.313 0.346 0.413 0.452 0.544 0.365 0.260 0.270 0.825 2.500 2.500 0.088 0.127 1.973 0.431 0.217 0.044 0.085 0.121 0.198 0.160 3.807 2.944 3.587 6.342  Flow Fan Power Operating Cost Fan Curve (m3/s) (kW) ($/year) Status 15.79 10.57 7,272 On 21.95 15.01 10,331 On 103.61 32.22 22,179 On 42.75 13.38 9,210 On 0.52 0.18 125 On 63.05 26.04 17,923 On 50.74 22.94 15,786 On 56.8 30.90 21,268 On 14.18 5.18 3,563 On 27.65 7.19 4,948 On 15.42 4.16 2,865 On 6.44 5.31 3,655 On 248.00 620.01 426,743 On 368.92 922.31 634,813 On 5.15 0.45 312 On 19.71 2.50 1,723 On 13.47 26.58 18,293 On 38.47 16.58 11,413 On 26.85 5.83 4,010 On 96.67 4.25 2,928 On 47.98 4.08 2,807 On 19.83 2.40 1,651 On 10.92 2.16 1,489 On 11.70 1.87 1,288 On 369.00 1,404.78 966,890 On 143.00 420.99 289,761 On 248.00 889.59 612,291 On 760.00 4,819.95 3,317,501 On 9,318.00 6,413,450  Fan Arrangement 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 2 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 1 in Parallel 2 in Parallel 2 in Parallel 2 in Parallel 3 in Parallel  Fan Location 7720L Booster Fan 7780L Booster Fan 7840L Booster Fan 7900L Booster Fan 7960L Booster Fan 8080L Booster Fan 8140L Booster Fan 8200L Booster Fan 8020L Booster Fan 8200L booster 8200L booster fan 3800L 5 Shaft fan 5000L Fans (2p) 5400L Fans 1900L 5 Shaft fan 6 shaft xc on 4800 L FAR 7200-7400 7840L fan 7840L to FAR/Esc. 7900L fan 7960L fan 8020L fan 8140L to ramp 8080L fan 2300L fans (2p) 1420L fans (2p) 2600L fans (2p) 11 Shaft Fans (3p)  206  Between years 2018 – 2020, the total fan power within the primary ventilation system is: PPrimary = 9,318 kW. For average 70% fan efficiency and for $0.055/kWh electricity charge, the primary ventilation system operating cost is: OCPrimary  (2018-2020)  = 9,318 kW/0.7 x 365  days/year x 24 hours/day x $0.055/kWh = $6,413,450/year. Auxiliary System – Flow Distribution, Fan Duties and Power Requirements For this time period to fulfill production requirements the mine would need to maintain 5 mining blocks active. Consequently, the combined fan power of the mine’s auxiliary ventilation system is: PAuxilairy = PProduction + PServices = 1,495 kW + 1,544 kW = 3,039 kW. Again, for an average 70% fan efficiency and $0.055/kWh electricity charge, the operating cost of the mine’s secondary ventilation system is: OCSecondary = 3,039 kW/0.70 x 365 days/year x 24 hours/day x $0.055/kWh = $2,091,700/year. For years (2018 – 2020) and traditional ventilation practice, the mine’s total fan power is: PTotal(2018-2020) = PPrimary = PAuxilairy = 9,318 kW + 3,039 kW = 12,357 kW. For average 70% fan efficiency and for $0.055/kWh electricity charge, the total operating is: OCTotal(2018-2020) = OCPrimary(2018-2020)  +  OCSecondary(2018-20220)  =  $6,413,450/year  +  $2,091,700/year  =  $8,505,150/year.  B.2  Solving and Balancing the Primary and Auxiliary Ventilation Systems of “Mine A” for Years 2011–2013, Years 2014-2017, Years 2018-2020 with Activity Based Air Volume Requirements  B.2.1 Ventilation Modelling for Years 2011–2013 (Activity Based Ventilation) Primary System – Flow Distribution, Fan Duties and Power Requirements Between years 2011-2013 and with activity based ventilation, the mine’s total intake volume (as calculated in section 7.2.2) is: QIntake-Activity(2011-2013) = 565 m3/s. For the same time period but based upon traditional ventilation practice the mine’s total intake air volume (as calculated in section 7.2.1) is: QIntake(2011-2013) = 775 m3/s. This indicates that with activity based ventilation, the mine’s intake air volume can be reduced by 210 m3/s, a 27% reduction. For activity based airflow requirements, the mine’s primary ventilation system was solved and balanced by means of ventilation simulation. The airflow distribution within the mine’s primary ventilation system is shown in Figure 48. 207  Within the primary system, the operating duty of the main fans is: •  1420 Level Intake Fans (2 in parallel): Q = 126m3/s, Ptotal = 435 Pa  •  2300 Level Intake Fans (2 in parallel): Q = 255m3/s, Ptotal = 983 Pa  •  2600 Level Intake Fans (2 in parallel): Q = 170m3/s, Ptotal = 1,770 Pa  •  No. 11 Shaft Exhaust Fans (3 in parallel): Q = 544m3/s, Ptotal = 3,181 Pa  The operating power of each individual fan (main/booster) that is part of the mine’s primary ventilation system is shown in Table 30. This table also shows that for years 2011-2013 and with activity based ventilation, the total fan power of the mine’s primary ventilation system is: PPrimary-Activity = 3,285 kW. For average 70% fans efficiency and for $0.055/kWh electricity charge, the operating cost of the mine’s primary ventilation system is: OCPrimary-Activity(2011-2013) = 3,285kW/0.7 x 365days/year x 24hours/day x $0.055/kWh = $2,261,020/year. Auxiliary System – Flow Distribution, Fan Duties and Power Requirements As previously mentioned, throughout this time-phase analysis the auxiliary ventilation setup within the mining blocks remains the same. Table 15 showed that the combined auxiliary fan power within one mining block is: PBlock-Activity = 127 kW. Between years 2011 - 2013, ore reserves will be extracted from 5 mining blocks. Consequently, the combined fan power within the active mining blocks is: PProduction-Activity = 127kW/block x 5 mining blocks = 635 kW. Again, assuming that 40% of the auxiliary service fans could be fully automated, the combined power for secondary services is: PServices-Activity = 1,186 kW. This shows that with activity based ventilation, the combined auxiliary fan power is: PAuxiliary-Activity = PProductionActivity  + PServices-Activity = 635 kW + 1,186 kW = 1,821 kW. For average 70% auxiliary fans  efficiency and $0.055/kWh electricity charge, the operating cost of the mine’s auxiliary ventilation system is: OCAuxiliary-Activity(2011-2013) = 1,821 kW/0.7 x 365 days x 24 hours/day x $0.055/kWh = $1,253,370/year. This shows that between years 2011 - 2013 and activity based ventilation, the mine’s total fan power is: PTotal-Activity(2011-2013) = PPrimary-Activity + PAuxiliary-Activity = 3,285 kW + 1,821 kW = 5,106 kW. For average 70% fan efficiency and $0.055/kWh electricity charge, the mine’s primary and auxiliary ventilation operating cost becomes: OCTotal-Activity(2011-2013) = OCPrimaryActivity(2011-2013)  +  OCAuxiliary-Activity(2011-2013)  =  $2,261,020/year  +  $1,253,370/year  =  $3,514,390/year.  208  Figure 48: Airflow distribution and fan operating duties in the primary system for years 2011 – 2013 - Activity based ventilation  209  Table 30: Fan Operating Duties and Power Requirements for Primary System (Years 2011 - 2013) – Activity based ventilation Junction From 383 39 380 420 487 7722 7726 7788 7784 7853 7845 7851 7908 7969 7905 8032 8090 7965 8025 8083 13 191 53 232 9 6 TOTAL  Junction To 384 40 381 421 485 7721 7727 7789 7783 7854 7844 7852 7909 7970 7906 8033 8091 7966 8026 8084 29 190 54 58 10 5  Pressure (kPa) 0.617 1.687 0.571 2.293 0.111 0.029 0.172 0.186 0.038 0.364 0.113 0.222 0.403 0.420 0.043 0.429 0.404 0.062 0.075 0.068 1.900 0.071 0.435 0.983 1.770 3.181  Flow Fan Power Operating Cost Fan Curve (m3/s) (kW) ($/year) Status 10.3 6.34 4,365 On 255.2 430.51 296,315 On 16.5 9.42 6,482 On 2.8 6.41 4,414 On 19.3 2.14 1,475 On 5.5 0.15 103 On 14.3 2.46 1,696 On 17.2 3.20 2,202 On 28.2 1.07 738 On 65.9 24.00 16,522 On 68.8 7.78 5,354 On 10.0 2.21 1,523 On 68.5 27.62 19,009 On 72.1 30.28 20,843 On 10.9 0.47 323 On 73.0 31.31 21,553 On 8.9 3.58 2,466 On 8.8 0.55 377 On 11.0 0.83 569 On 33.0 2.24 1,542 On 170.0 322.99 222,311 On 6.2 0.44 305 On 126.0 54.81 37,724 On 255.0 250.68 172,539 On 170.0 300.89 207,100 On 554.0 1762.30 1,212,963 On 3,285.00  2,261,020  Fan Arrangement 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 2 in Parallel 2 in Parallel 2 in Parallel 3 in Parallel  Fan Location 3800 LEVEL 5 SHAFT 5400L Fans 3800L Booster Fan 4200L Booster Fan 4800L Booster Fan 7660 FAR booster fan 7720L Booster fan 7780L Booster Fan 7660 FAR Booster 7840L Booster fan 7660 FAR Booster 7840 Esc. Booster 7900L Booster fan 7960L Booster Fan 7900L booster 8020L Booster fan 8080L Booster Fan 7960L booster 8020L booster 8080L booster 5000L Fans (2p) 1900 LEVEL 5 SHAFT 1420L Fans (2p) 2300L Fans (2p) 2600L Fans (2p) 11 Shaft Fans (3p)  210  For the same time period but traditional ventilation practice (see section B.1.1) the mine’s total fan power and operating cost was calculated at: PTotal(2011-2013) = 10,512 kW and OCTotal(2011-2013) = $7,235,300/year, respectively. The above calculations indicate that with activity based ventilation, a reduction in the mine’s intake air volume of 210 m3/s (27%) can generate power consumption savings of: PSTotal(2010-2013) = (10,512 kW – 5,106 kW)/0.7 x 365 days/year x 24 hours/day = 67,652.2 MWh/year or $3,720,910/year, a 51% reduction.  B.2.2 Ventilation Modelling for Years 2014–2017 (Activity Based Ventilation) Primary System – Flow Distribution, Fan Duties and Power Requirements Between years 2014-2017 and with activity based ventilation practice, the mine’s total intake volume was calculated at (see section 7.2.2): QIntake-Activity(2014-2017) = 638 m3/s. For the same time period but traditional ventilation practice the mine’s total intake air volume (see section 7.2.1) is: QIntake(2014-2017) = 890 m3/s. This indicates that with activity based ventilation, the mine’s intake air volume can be reduced by 252 m3/s, a 28% reduction. The airflow distribution within the mine’s primary ventilation system is shown in Figure 49. Within the primary system the fan operating duties are: •  1420 Level Intake Fans (2 in parallel): Q = 138 m3/s, Ptotal = 1,313 Pa  •  2300 Level Intake Fans (2 in parallel): Q = 290 m3/s, Ptotal = 2,626 Pa  •  2600 Level Intake Fans (2 in parallel): Q = 196 m3/s, Ptotal = 2,507 Pa  •  No. 11 Shaft Exhaust Fans (3 in parallel): Q = 613 m3/s, Ptotal = 3,860 Pa  The operating power of each individual fan (main/booster) that is part of the mine’s primary ventilation system is shown in Table 31. This table also shows that for years 2014-2017 and with activity based ventilation, the total fan power of the mine’s primary ventilation system is: PPrimary-Activity = 4,933 kW. For average 70% fans efficiency and $0.055/kWh electricity charge, the operating cost of the mine’s primary ventilation system is: OCPrimary-Activity(2014-2017) = 4,933kW/0.7 x 365days/year x 24hours/day x $0.055/kWh = $3,395,310/year.  211  Figure 49: Airflow distribution and fan operating duties in the primary system for years 2014 - 2017 - Activity based ventilation  212  Table 31: Fan operating duties and power requirements for primary system (years 2014 - 2017) – Activity based ventilation Junction From 7726 7788 7853 7908 7969 8090 8152 8210 8032 8205 8208 383 13 39 191 487 36 7844 7851 7906 7965 8025 8146 8083 53 232 9 6 TOTAL  Junction To 7727 7789 7854 7909 7970 8091 8153 8211 8033 8206 8209 384 29 40 190 485 35 7843 7852 7907 7966 8026 8147 8084 54 58 10 5  Pressure (kPa) 0.669 0.684 0.311 0.313 0.346 0.413 0.452 0.544 0.365 0.260 0.270 1.500 1.600 1.600 0.100 0.127 1.973 0.431 0.217 0.044 0.085 0.121 0.198 0.160 1.313 2.626 2.507 3.860  Flow Fan Power Operating Cost Fan Curve (m3/s) (kW) ($/year) Status 29.61 19.81 13,632 On 21.76 14.88 10,245 On 100.79 31.35 21,575 On 38.91 12.18 8,383 On 2.33 0.80 554 On 62.86 25.96 17,870 On 50.73 22.93 15,783 On 56.79 30.89 21,264 On 13.81 5.04 3,469 On 27.62 7.18 4,943 On 15.42 4.16 2,866 On 14.29 21.43 14,753 On 196 313.61 215,851 On 289.93 463.88 319,284 On 7.14 0.71 492 On 20.17 2.56 1,763 On 51.19 101.00 69,515 On 50.90 21.94 15,100 On 28.31 6.14 4,229 On 96.85 4.26 2,933 On 48.75 4.14 2,852 On 19.39 2.35 1,615 On 10.96 2.17 1,493 On 11.91 1.91 1,312 On 138.00 181.19 124,713 On 289.99 761.51 524,139 On 196.00 491.38 338,211 On 615.99 2,377.74 1,636,563 On 4,933.00 3,395,310  Fan Arrangement 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 2 in Parallel 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 1 in Series 2 in Parallel 2 in Parallel 2 in Parallel 3 in Parallel  Fan Location 7720L Booster Fan 7780L Booster Fan 7840L Booster Fan 7900L Booster Fan 7960L Booster Fan 8080L Booster Fan 8140L Booster Fan 8200L Booster Fan 8020L Booster Fan 8200L booster 8200L booster fan 3800L 5 Shaft fan 5000L Fans (2p) 5400L Fans 1900L 5 Shaft fan 6 shaft x-cut on 4800 L FAR 7200-7400 7840L fan 7840L to FAR/Esc. 7900L fan 7960L fan 8020L fan 8140L to ramp 8080L fan 1420L Fans (2p) 2300L Fans (2p) 2600L Fans (2p) 11 Shaft fans (3p)  213  Auxiliary System – Flow Distribution, Fan Duties and Power Requirements Between years 2014-2017, to satisfy production requirements 6 mining block would need to be maintained active. Consequently, the combined fan power within the active mining blocks is: PProduction-Activity = 127 kW/block x 6 mining blocks = 762 kW. Again, assuming that 40% of the auxiliary service fans could be controlled, the combined power for secondary services would be: PServices-Activity = 1,186 kW. This shows that with activity based ventilation, the total fan power of the mine’s auxiliary ventilation system is: PAuxiliary-Activity = PProduction-Activity + PServices-Activity = 762 kW + 1,186 kW = 1,948 kW. For average 70% fans efficiency and $0.055/kWh electricity charge, the operating cost of the mine’s auxiliary ventilation system would be: OCAuxiliary-Activity(2014-2017) = 1,948 kW/0.7 x 365 days x 24 hours/day x $0.055/kWh = $1,340,780/year. This shows that between years 2014 - 2017 and activity based ventilation, the mine’s total fan power is: PTotal-Activity(2014-2017) = PPrimary-Activity + PAuxiliary-Activity = 4,933 kW + 1,948 kW = 6,881 kW. Again, for average 70% fans efficiency and $0.055/kWh electricity charge, the mine’s primary and auxiliary ventilation operating cost becomes: OCTotal-Activity(2014-2017) = OCPrimary-Activity(2014-2017) + OCAuxiliary-Activity(2014-2017) = $3,395,310/year + $1,340,780/year = $4,736,090/year. For the same time period but with traditional ventilation (see section B.1.2) the mine’s total fan power and operating cost was calculated at: PTotal(2014-2017) = 17,916 kW and OCTotal(20112013)  = $12,331,330/year, respectively. This indicates that with activity based ventilation, a  reduction in the mine’s intake air volume of 252 m3/s, a 28% reduction, can generate combined power consumption savings within the primary and auxiliary systems in the range of: PSTotal(2013-2017) = (17,916 kW – 6,881 kW)/0.7 x 365 days/year x 24 hours/day = 138,095.1 MWh/year or $7,595,240/year, a 61% reduction.  B.2.3 Ventilation Modelling for Years 2018–2020 (Activity Based Ventilation) Primary System – Flow Distribution, Fan Duties and Power Requirements For years 2018-2020 and with activity based ventilation, the mine’s total intake volume as determined in section 7.2.2 is: QIntake-Activity(2018-2020) = 565 m3/s. For the same time period but with traditional ventilation, the mine’s intake air volume was calculated at: QIntake(2018-2020) =  214  775 m3/s. This indicates that with activity based ventilation, between years 2018 – 2020, the mine’s intake air volume can be reduced by 210 m3/s, a 27% reduction. The air volume requirements in this scenario are similar to the air volume requirements for years 2011 – 2013. However, in this scenario the mine resistance has a higher value due to increased mining depth. Between years 2018 – 2020, the airflow distribution within the primary ventilation system is shown in Figure 50. Within the primary system, the primary fan operating duties are: •  1420 Level Intake Fans (2 in parallel): Q = 133 m3/s, Ptotal = 787 Pa  •  2300 Level Intake Fans (2 in parallel): Q = 253 m3/s, Ptotal = 1,643 Pa  •  2600 Level Intake Fans (2 in parallel): Q = 169 m3/s, Ptotal = 1,261 Pa  •  No. 11 Shaft Exhaust Fans (3 in parallel): Q = 553 m3/s, Ptotal = 3,371 Pa  The operating power of each individual fan (main/booster) that is part of the mine’s primary ventilation system is shown in Table 32. This table also shows that for years 2018 - 2020 and activity based ventilation, the total fan power of the mine’s primary ventilation system is: PPrimary-Activity = 3,547 kW. For average 70% fans efficiency and $0.055/kWh electricity charge, the operating cost of the mine’s primary ventilation system is: OCPrimary-Activity(2018-2020) = 3,547kW/0.7 x 365days/year x 24hours/day x $0.055/kWh = $2,441,350/year. Auxiliary System – Flow Distribution, Fan Duties and Power Requirements Between years 2018 - 2020 ore reserves will be extracted from 5 mining blocks. Consequently, the total fan power of the mine’s auxiliary ventilation system is similar to 2011 - 2013 time period, which was calculated at: PAuxiliary-Activity = PProduction-Activity + PServicesActivity  = 635 kW + 1,186 kW = 1,821 kW. With activity based ventilation, the operating cost  of the mine’s auxiliary ventilation system is: OCAuxiliary-Activity(2014-2017) = 1,821 kW/0.7 x 365 days x 24 hours/day x $0.055/kWh = $1,253,370/year. This shows that between years 2018 - 2020 and with activity based ventilation, the mine’s total fan power is: PTotal-Activity(2018-2020) = PPrimary-Activity + PAuxiliary-Activity = 3,547 kW + 1,821 kW = 5,368 kW. Again, for average 70% fans efficiency and $0.055/kWh electricity charge, the mine’s primary and auxiliary ventilation operating cost becomes: OCTotal-Activity(2014-2017) = OCPrimary-Activity(2018-2020) + OCAuxiliary-Activity(2018-2020) = $2,441,350/year + $1,253,370/year = $3,694,720/year. 215  Figure 50: Airflow distribution and fan operating duties in the primary system for years 2018 - 2020 - Activity based ventilation  216  Table 32: Fan operating duties and power requirements for primary system (years 2018 - 2020) – Activity based ventilation Junction From 7726 7788 7853 7908 7969 8090 8152 8210 8032 8205 8208 383 13 39 191 487 36 7844 7851 7906 7965 8025 8146 8083 53 232 9 6 TOTAL  Junction To 7727 7789 7854 7909 7970 8091 8153 8211 8033 8206 8209 384 29 40 190 485 35 7843 7852 7907 7966 8026 8147 8084 54 58 10 5  Pressure (kPa) 0.669 0.684 0.311 0.313 0.346 0.413 0.452 0.544 0.365 0.260 0.270 0.825 1.500 1.300 0.088 0.127 1.973 0.431 0.217 0.044 0.085 0.121 0.198 0.160 0.787 1.643 1.261 3.371  Flow Fan Power Operating Cost Fan Curve Fan (m3/s) (kW) ($/year) Status Arrangement 40.60 27.16 18,694 On 1 in Parallel 26.21 17.93 12,341 On 1 in Parallel 98.88 30.75 21,165 On 1 in Parallel 37.05 11.60 7,981 On 1 in Parallel 3.06 1.06 730 On 1 in Parallel 62.95 26.00 17,894 On 1 in Parallel 50.65 22.89 15,758 On 1 in Parallel 56.79 30.89 21,262 On 1 in Parallel 13.27 4.84 3,333 On 1 in Parallel 27.65 7.19 4,948 On 1 in Parallel 15.42 4.16 2,866 On 1 in Parallel 11.00 9.07 6,243 On 1 in Parallel 169.01 253.52 174,495 On 2 in Parallel 252.92 328.80 226,310 On 1 in Parallel 6.88 0.61 417 On 1 in Parallel 20.24 2.57 1,769 On 1 in Parallel 62.14 122.60 84,381 On 1 in Parallel 61.02 26.30 18,101 On 1 in Parallel 30.23 6.56 4,515 On 1 in Parallel 96.98 4.27 2,937 On 1 in Parallel 48.97 4.16 2,865 On 1 in Parallel 19.72 2.39 1,642 On 1 in Parallel 10.96 2.17 1,493 On 1 in Parallel 11.67 1.87 1,285 On 1 in Parallel 133.00 104.67 72,044 On 2 in Parallel 253.01 415.69 286,114 On 2 in Parallel 169.01 213.13 146,692 On 2 in Parallel 553.02 1,864.24 1,283,129 On 3 in Parallel 3,547.00 2,441,350  Fan Location 7720L Booster Fan 7780L Booster Fan 7840L Booster Fan 7900L Booster Fan 7960L Booster Fan 8080L Booster Fan 8140L Booster Fan 8200L Booster Fan 8020L Booster Fan 8200L booster 8200L booster fan 3800L 5 Shaft fan 5000L Fans (2p) 5400L Fans 1900L 5 Shaft fan 6 shaft x-cut on 4800 L FAR 7200-7400 7840L fan 7840L to FAR/Esc. 7900L fan 7960L fan 8020L fan 8140L to ramp 8080L fan 1420L Fans (2p) 2300L Fans (2p) 2600L Fans (2p) 11 Shaft Fans (3p)  217  For the same time period but traditional ventilation practice (see section B.1.3), the mine’s total fan power and operating cost was calculated at: PTotal(2014-2017) = 12,357 kW and OCTotal(2011-2013) = $8,505,150/year, respectively. This indicates that between years 2018 – 2020 and with activity based ventilation, a reduction in the mine’s intake air volume of 210 m3/s, (27% reduction) could generate power consumption savings within the primary and auxiliary systems in the range of: PSTotal(2013-2017) = (12,357 kW – 5,368 kW)/0.7 x 365 days/year x 24 hours/day = 87,462.3 MWh/year or $4,810,430/year, a 56% reduction in ventilation operating costs.  218  

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