Carefully engineered mine ventilation is critical to the safe operation of underground longwall mines. Currently, there are several options for simulation of mine ventilation. This research was conducted to rapidly simulate an underground longwall mine, especially for the use of tracer gas in an emergency situation. In an emergency situation, limited information about the state of mine ventilation system is known, and it is difficult to make informed decisions about safety of the mine for rescue personnel. With careful planning, tracer gases can be used to remotely ascertain changes in the ventilation system. In the meantime, simulation of the tracer gas can be conducted to understand the airflow behavior for improvements during normal operation. Better informed decisions can be made with the help of both tracer gas technique and different modeling approaches. This research was made up of two main parts. One was a field study conducted in an underground longwall mine in the western U.S. The other one was a simulation of the underground longwall mine with different approaches, such as network modeling and Computational Fluid Dynamics (CFD) models. Networking modeling is the most prevalent modeling technique in the mining industry. However, a gob area, which is a void zone filled with broken rocks after the longwall mining, cannot be simulated in an accurate way with networking modeling. CFD is a powerful tool for modeling different kinds of flows under various situations. However, it requires a significant time investment for the expert user as well as considerable computing power. To take advantage of both network modeling and CFD, the hybrid approach, which is a combination of network modeling and CFD was established. Since tracer gas was released and collected in the field study, the tracer gas concentration profile was separately simulated in network modeling, CFD model, and hybrid model in this study. The simulated results of airflow and tracer gas flow were analyzed and compared with the experimental results from the field study. Two commercial network modeling software packages were analyzed in this study. One of the network modeling software also has the capability to couple with CFD. A two-dimensional (2D) CFD model without gob was built to first analyze the accuracy of CFD. More 2D CFD models with gob were generated to determine how much detail was necessary for the gob model. Several three-dimensional (3D) CFD models with gob were then created. A mesh independence study and a sensitivity study for the porosity and permeability values were created to determine the optimal mesh size, porosity and permeability values for the 3D CFD model, and steady-state simulation and transient simulations were conducted in the 3D CFD models. In the steady-state simulation, a comparison was made between the 3D CFD models with and without taking the diffusivity of SF6 in air into account. Finally, the different simulation techniques were compared to measured field data, and assessed to determine if the hybrid approach was considerably simpler, while also providing results superior to a simple network model. / Master of Science

A thesis submitted to the Faculty of Engineering and the Built Environment, University of the Witwatersrand, Johannesburg, in fulfilment of the requirements for the degree of Doctor of Philosophy. Johannesburg, 2016 / The blind headings created in room and pillar mining are known to be the high risk areas of the coal mine, since this is where the coal production is actually taking place and hence the liberation of maximum quantity of methane. The ventilation of this region called the localized ventilation is carried out using auxiliary ventilation devices. This ventilation may be planned and be the subject of mine standards, but it is not very well understood and implementation on a day to day basis is usually left to the first level of supervisory staff. Majority of the methane explosions have been found to occur in these working areas and blind headings. The correct use of auxiliary ventilation devices can only be carried out once the effect of the system variables associated with each device is very well understood and can be calculated mathematically. Presently, no mathematical models or empirical formulas exist to estimate the effect of the associated system variables on the flow rates close to the face of the heading. The extent of ventilation of a heading ventilated without the use of any auxiliary device is not clear. Furthermore, to design additional engineering solutions, the flow patterns inside these heading ventilated with the auxiliary ventilation devices needs to be understood. The study of the face ventilation systems and the effect of the system variables associated system with each auxiliary ventilation device can be carried out experimentally, but doing a large number of experiments underground is very difficult as it disturbs the mine production cycles. Furthermore, studying the flow patterns experimentally is even more cumbersome, and can only be done to some extent using smoke or tracer gas. Therefore, Computational Fluid Dynamic‟s (CFD) advanced numerical code ANSYS Fluent was used to study the effect of a number of system variables associated with the face ventilation systems used in blind headings. As part of the procedure, the CFD model used was validated using four validation studies, in which the numerical results were compared with the actual experimental results. The numerical results differed to a maximum of 10% for all the experimental results. The system variables associated with ventilation of a heading, without the use of any auxiliary device, with the use of Line Brattice (LB) and fan with duct were selected. A range of values was chosen for each variable, and scenarios were created using every possible combination of these variables. All the scenarios were simulated in Ansys Fluent, the air flow rates, air velocities, velocity vectors, and velocity contours were calculated and drawn at different locations inside the heading. The effect of each system variable was found using a comparative analysis. The results were represented in simple user-friendly form and can be used to estimate the air flows at the exit of the LB and face of the heading for various settings of the LB and fan and duct face ventilation systems. The analysis of the ventilation of a heading without the use of LB shows that a maximum penetration depth is found with the Last Through Road (LTR) velocity of 1.35m/s. The flow rates and the maximum axial velocities increase with the increase in the LTR velocity up to a depth of 10m (maximum air flowing into a heading of 1.26m3/s and 1.58m3/s is found for the 3m and 4m high heading using 2m/s LTR velocity). For the LB ventilation system the LTR velocities, heading height, length of the LB in the LTR and heading, angle of the LB in LTR, and distance of the LB to the wall of the heading (side wall) were varied to identify clearly the effect of these control variables, on the flow rate at the exit of the LB, and close to the face of the heading. The flow rate at the exit of the LB is found to be proportional to the product of the distance of the LB to the wall in the LTR and heading. The flow rate at the exit of the LB, face of the heading, and inside the heading is found proportional to the LTR velocity and height of the heading. It is found that a minimum length of LB is associated with each distance of the LB to the wall in the heading, to maximize the delivery of air close to the face of the heading. This length is found to be equal to 15m for 1m LB to wall distance, and 10m for 0.5m LB to wall distance. Mathematical models were developed to estimate the effect of each studied system variables on the flow rates at the exit of the LB and close to face of the heading. For the fan and duct systems the length, diameter, and the fan design flow rates were varied. It is found that for a force fan duct system only a maximum of 50% of the total air that reaches the face is fresh and the remaining 50% is recirculated air. The flow rate with the exhaust fan system is found to be much lower than the force fan duct system. It increases with the reduction in duct mouth to heading face distance, and increase in duct diameter. Mathematical models are developed to calculate the flow rates at the face of the heading using the effect of each studied system variable. The research reveals that the ANSYS numerical code is an appropriate tool to evaluate the face ventilation of a heading in a three dimensional environment using full scale models. The South African coal mining industry can benefit from the outcomes of this study, specially the mathematical models, in a number of ways. Ventilation engineers can now estimate the flow rates close to the face of the heading for different practical mining scenarios and ensure sufficient ventilation by using the appropriate auxiliary ventilation settings. The results can easily be developed into training aids using easy to use excel spread sheets to ensure that mineworkers at the coal face have a better understanding of the working of the auxiliary ventilation devices. It can also serve Academia as part of the curriculum to teach the future mining engineers how the different variables associated with the auxiliary ventilation system affect the ventilation in a heading. The research therefore, has the potential to provide a significant step toward, understanding airflow rates delivered by the auxiliary devices close to the face of the heading and the air flow patterns inside the heading as a basis for improving the working environment for underground mineworkers. / MT2017

A comprehensive experimental study has been conducted using Particle Image Velocimetry (PIV) for a wide array of ventilation schemes and mining configurations for the purpose of examining ventilation characteristics in dead end mine headings. Flow behaviors in two basic mining sequences of box and slab cuts for 30 ft and 60 ft deep cuts were studied. The present thesis discusses the effect for various geometric and flow parameters including the variation of inlet flow velocities, entry heights, face zone widths and curtain widths on the flow behavior. The Reynolds number Re considered for this study ranges from 1 105 to 3 106 based on curtain width and exit velocity. The variation of the face zone and the curtain widths considerably affected the flow behavior, resulting in recirculation regions in the face area for critical combinations. Jet spreading angles and virtual origins have been calculated for the different geometries showing that an optimum range of face and curtain widths exists. A detailed discussion employing various scenarios for exhaust ventilation systems has also been made. Full-size measurements and comparison of the experimental data with numerical simulations is presented. Implementation of machine-mounted scrubbers in the blowing system are discussed for different values of the ventilation ratios (Qs/Qin) ranging from 14% to 53%. The scrubber system, typically introduced for dust collection, is also shown to be a useful tool in providing adequate ventilation to the immediate face area.

No description available.

No description available.

Thesis (M.E.) -- University of Adelaide, 1956. / Typewritten copy.

Thesis (M.S.)--University of Nevada, Reno, 2007. / "August, 2007." Includes foldout illustrations. Includes bibliographical references (leaves 79-84). Library also has microfilm. Ann Arbor, Mich. : ProQuest Information and Learning Company, [2008]. 1 microfilm reel ; 35 mm. Online version available on the World Wide Web.

Increasing energy costs and the need to conserve energy compounded with low mineral prices have prompted some Canadian mines especially potash producers, to examine their operations and identify potential saving methods. Re-using or recirculating a fraction of ventilating air may enable these mines to reduce winter heating costs. Gas and dust concentrations were monitored in the intake and exhaust airways to assess the potential for recirculating exhaust air. The results indicate that the mine pollutant concentrations in potash mines are low and stable. Trial recirculation experiments returning 20 - 47% exhaust air into the fresh air airway did not cause significant increases in mixed intake pollutant levels. Two types of recirculation systems - namely variable and fixed quantity - are developed. Detailed designs of recirculation systems for Central Canada Potash of Noranda Minerals Inc. and Rocanville Division, Potash Corporation of Saskatchewan are discussed and recommendations made for the selection and positioning of on-line monitoring, control and telemetry systems. A controlled recirculation system conceptual design for the H-W mine is given. The economic payback periods for systems proposed for Rocanville Division and CCP are 2 and 3 years respectively. Recirculation percentages of 30%, 64.4% and 23% are feasible for CCP, Rocanville Division and the H-W mine. The recirculation percentages for the proposed systems were determined using Air Quality Index criteria. Dust deposition studies conducted at CCP in return airways indicate that 65% of dust by weight is deposited within a distance of 550 metres from the face. In terms of dust and other contaminant conditions in the return airways, it can be concluded that there is potential for use of recirculation in the face area. Guidelines for recirculation systems in gassy and dusty mines are developed. The main features for these recirculation system design guidelines are safety, economic gain, and system performance. The author's attribution to ventilation is in the use of controlled recirculation to reduce winter heating costs and increase underground airflow, also the guidelines developed for recirculation in gassy and dusty mines. The overall conclusion is that controlled recirculation is a practical method of reducing winter heating cost and/or increasing mine airflows. The financial potential and technology to implement a working system exist. / Applied Science, Faculty of / Mining Engineering, Keevil Institute of / Graduate

Following an unexpected event in an underground mine, it is important to know the state of the mine immediately to manage the situation effectively. Particularly when part or the whole mine is inaccessible, remotely and quickly ascertaining the ventilation status is one of the pieces of essential information that can help mine personnel and rescue teams make decisions. This study developed a methodology that uses tracer gas techniques and CFD modeling to analyze underground mine ventilation system status remotely. After an unanticipated event that has damaged ventilation controls, the first step of the methodology is to assess and estimate the level of the damage and the possible ventilation changes based on the available information. Then CFD models will be built to model the normal ventilation status before the event, as well as possible ventilation damage scenarios. At the same time, tracer gas tests will be designed and performed on-site. Tracer gas will be released at a designated location with constant or transient release techniques. Gas samples will be collected at other locations and analyzed using Gas Chromatography (GC). Finally, through comparing the CFD simulated results and the tracer on-site test results, the general characterization of the ventilation system can be determined. A review of CFD applications in mining engineering is provided in the beginning of this dissertation. The basic principles of CFD are reviewed and six turbulence models commonly used are discussed with some examples of their application and guidelines on choosing an appropriate turbulence model. General modeling procedures are also provided with particular emphasis on conducting a mesh independence study and different validation methods, further improving the accuracy of a model. CFD applications in mining engineering research and design areas are reviewed, which illustrate the success of CFD and highlight challenging issues. Experiments were conducted both in the laboratory and on-site. These experiments showed that the developed methodology is feasible for characterizing underground ventilation systems remotely. Limitations of the study are also addressed. For example, the CFD model requires detailed ventilation survey data for an accurate CFD modeling and takes much longer time compared to network modeling. Some common problems encountered when using tracer gases in underground mines are discussed based on previously completed laboratory and field experiments, which include tracer release methods, sampling and analysis techniques. Additionally, the use of CFD to optimize the design of tracer gas experiments is also presented. Finally, guidelines and recommendations are provided on the use of tracer gases in the characterization of underground mine ventilation networks. / Ph. D.

I underjordiska gruvor behöver stora mängder luft tillföras. När utetemperaturen sjunker under 0 °C leder detta till stort uppvärmningsbehov för att fukten i tilluftschaktet inte ska frysa, vilket inte bara är kostsamt, utan även negativt ur ett hållbarhetsperspektiv. Vanligtvis återvinns ca 35 % av det årliga energibehovet för uppvärmningen. Med en roterande värmeväxlare anpassad för gruvventilation är det möjligt att återvinna hela energibehovet för att värma tilluften över 0 °C för att förhindra isbildning i luftschaktet. Syftet med projektet var att undersöka och beräkna värme- (och energi) överföring i en roterande värmeväxlare som funktion av luftflödet, effekten för att värma tilluften och dimension på värmeväxlarens rotordiameter.   Kontaktarean på ackumulatormaterialet av en pilotanläggning har beräknats för att undersöka hur stor värmeöverförande yta som förekommer och hur denna yta påverkas vid olika dimensioner på materialet. Teoretisk värmeöverföring och värmeöverföringen av pilotanläggningen vid kallt klimat har beräknats. Dessa resultat har sedan använts för att dimensionera värmeväxlarens rotordiameter i full skala.    Den totala värmeöverföringsytan vid med pilotanläggningens dimensioner beräknades till 4086 m2. Den teoretiska värmeöverföringen beräknades till 214 kW. Vid uppskalning till ett vanligt förekommande luftflöde i full skala (150 m3/s) beräknades värmeöverföringen till 2988 kW, kommer en rotordiameter på 11 m krävas. Eftersom beräkningarna inte tar hänsyn till praktiska förluster, är resultatet en grov uppskattning. Möjligheten att jämföra resultatet med andra befintliga anläggning är begränsad, eftersom denna typ av roterande värmeväxlare ej förekommer i gruvventilation idag. / In underground mines, large amounts of fresh air need to be supplied. When the outdoor temperature drops below 0 °C, the supplied outdoor air must be heated to prevent the moisture in the supply air shaft from freezing. This is not only expensive, but also negative from a sustainability perspective. Today usually about 35 % of the annual energy demand of the heating is recycled. With a rotary heat exchanger adapted for mine ventilation, it is possible to recover all the energy required to heat the supply air to at least 0 °C. The purpose of this project was to investigate and calculate the heat (and energy) transfer in a rotary heat exchanger as a function of the air flow, the required effect to heat the supply air, and the dimension of the rotor diameter of the heat exchanger.    The contact area of the accumulator material of a pilot plant was calculated to investigate the heat transfer surface and how this surface is affected by different dimensions of the material. The theoretical heat transfer and the heat transfer of the pilot plant in cold climate were also determined. These results were used to provide an estimate of the rotor diameter of the heat exchanger required in full scale.   The total heat transfer surface with unchanged dimensions was calculated to 4086 m2. The theoretical heat transfer was calculated to 241 kW. When upscaling to a common air flow at full scale (150 m3/s) the heat transfer was calculated to 2988 kW, a rotor diameter of 11 m will be required. Since the calculations do not consider (energy) losses that will occur in practise, this estimate is conservative. The possibility to compare results with a similar exchanger is limited, since this type of rotary heat exchanger is not used in mine ventilation today.