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By D. F. Caicedo, I. A. Brum, L. A. Betancourt

Ferricyanide ()()*is one of the most stable cyanometallic complexes present in the gold mining effluents. This cyanocomplex is hard to degrade by natural attenuation and generates a negative impact on aquatic environments. Although free cyanide (CN-) can be obtained by acidifying the solution, the CN- is lethal for all forms of life. The oxidation of ()()* in a typical photocatalitic system was evaluated with the addition of H2O2. To verify the degradation, chemical parameters, such as free cyanide, the formation of ammonia, nitrate, and total iron were analysed at the end of the process. The degradations obtained were up to 70% in a heterogeneous photocatalysis system with TiO2, and a low rate of dissociation was observed by photolysis. The best result was achieved with the H2O2 and TiO2 photocatalytic system, and the stoichiometric concentration was about 2.5 times less than the peroxide used in the real gold mining industry, reaching 83% degradation. The photocatalytic process obtained less toxic byproducts than the original synthetic ferricyanide used as mining wastewater. Keywords: Cyanide, heterogeneous photocatalysis, titanium dioxide, UV radiation, gold wastewater

Jan 1, 2020

By Rimas Veitas, James Panton, Erich Steinlechner

"Ductile Iron Piles (DIPs) are a proven, viable alternative to conventional piles, and for the past five years have been successfully used throughout New England. The installation process of DIPs lends itself to both large and small projects. Because of their installation efficiency they provide a substantial cost savings over drilled micro piles. Several New England projects (both end bearing and friction) will be evaluated, focusing on the installation and design procedures.BackgroundThe driven ductile cast iron pile (DIP) is a simple, fast and highly effective pile system. Over the last 30 years over 10 million feet of this pile system has been installed worldwide predominantly in Europe. DIPs are a prefabricated driven pile system utilizing high strength ductile iron pipes which are manufactured using a spun-cast process having outside diameters of 4 5/8” and 6 5/8 inches. The standard pile lengths are 16.4 feet long. The pipes are manufactured with a tapered socket with an internal shoulder for full engagement at the top and a tapered spigot at the bottom. The individual pile sections can be connected with this Plug and Drive® connection to drive a pile of any length.(see figure 1)The DIP system’s inherent advantage is the simplicity of the installation process. The piles are driven to refusal or to a required penetration depth with the use of a high-impact highfrequency hydraulic hammer (ie. Breaker hammer). In most applications the hydraulic hammer is mounted on an appropriately sized excavator (see Photo 2). At the completion of driving the pile, the pile is cut off (see photo 5) and the remaining section of pile is fitted with the appropriate driving shoe and serves as the lead section for the next pile. Being able to reuse the cut off section, yields minimal if any pile shaft waste. The pile is then fitted with an appropriate pile cap plate to receive the load from the superstructure."

Jan 1, 2017

By D J. Williams

Tailings management is typically constrained by many factors that are taken as given. These include the tailings production rate that must be accommodated, the desire to deliver the tailings cheaply and easily as a slurry using centrifugal pumps, and a tailings footprint limited by the desire to minimise capital costs leading to a rapid tailings rate of rise and a low settled density. Other constraints include wall raises being delayed due to limited capital, inadequate tailings water and rainfall run-off management, and the net present value (NPV) approach to capital expenditure that puts off expenditure at the expense of performance. All of these constraints result in the tailings occupying much more volume than is necessary, being difficult to rehabilitate, and not being capable of supporting a beneficial post-mining land use or function. Ultimately, such constrained tailings storage facilities will assume very large dimensions, but each will typically be developed as a series of smaller storages that are filled rapidly and added to progressively. The optimal stage to efficiently recover process water is in the process plant, but this is limited by the desire to use centrifugal pumps rather than the more costly positive displacement pumps, which also require greater management of tailings deposition. Process water can also be recovered from the tailings storage facility (TSF) decant. Tailings water that is not recovered from the plant or the decant will be lost to entrainment within the tailings, seepage to the foundation or through the wall, or evaporation. However, tailings water is not always suitable for reuse in the plant, particularly when it becomes contaminated by the processing, and it may simply be stored on the TSF. This will exacerbate the seepage of contaminated water to the environment, and will limit the densification of the tailings that would otherwise occur due to desiccation of the tailings on exposure. If a larger tailings footprint were developed early, cycling of tailings deposition in thin lifts between a number of cells could occur, maintaining the dormant cells dry to enhance desiccation and densification, and ultimate rehabilitation. The end result would be a much reduced stored volume of tailings, the potential for upstream raising, and ease of rehabilitation to a high level. The higher initial capital cost would be overtaken by reduced ongoing capital, operating and closure costs.CITATION:Williams, D J, ?2014. An alternative whole-of-life approach to tailings management, in Proceedings Life-of-Mine 2014 , pp 285–298 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 16, 2014

By Thomas E. Lobb, Harry C. Verakis

Blasting is a primary means of extracting minerals and ores at surface mining operations. The domestic consumption of explosives and blasting agents during the year 2001 was about 5.25 billion pounds. Flyrock is always a major concern for the blaster. Flyrock from surface blasting operations has caused serious injury and death to employees and other persons. Injuries due to flyrock and the lack of blast area security accounted for over two-thirds of all blasting-related injuries in surface coal, metal, and nonmetal mines during the period 1978-2002. Selected accidents due to flyrock and lack of blast area security in surface mining are presented in this paper. Incidents related to construction blasting are also described. Techniques to mitigate blasting accidents are discussed. These include proper blast design, driller-blaster communication, inspection prior to loading and firing the blast, removing employees from the blast area, controlling access to the blast area, and using a blasting shelter. An experienced driller could detect potential problem areas such as voids, mud seams, incompetent rocks, and other irregularities by observing the progress of drilling. The drill log should include the details of any unusual or exceptional circumstances noticed during drilling. A blaster may need to alter the loading configuration to alleviate potential problems. Basic blast design is sometimes taken for granted and assumed to be proper for the conditions encountered, but one size does not fit all. It is known from the physics of blasting that the explosive energy takes the path of least resistance. The path of least resistance could generate flyrock, depending on the blast site conditions. A combination of ‘borehole tracking’ and ‘laser profiling’ can assist in improving the design of a blast. The blaster can use these tools to adjust borehole loading to match site conditions.

Jan 1, 2004

By Jr. Wallace

This paper describes an analysis performed to determine the cool-down time of drifts in the prospective commercial nuclear waste repository in volcanic tuff at Yucca mountain, Nevada. Two emplacement methods are currently being considered at Yucca mountain. The reference configuration stores the high-level waste in short vertical boreholes, and the alternative configuration stores the waste in long horizontal bore-holes. After emplacement operations for each emplacement drift are complete, the drift will be closed off from continuous ventilation. Given time, the radioactive decay of high-level waste will increase the rock temperature around access and emplacement drifts. Air cooling may be required to re-enter these drifts. The paper discusses the analysis procedure and computer programs used to evaluate the time to cool a drift to acceptable environmental conditions 50 years after emplacement of high-level waste. Key assumptions to the analysis are presented. The results of the study show that cooling either vertical or horizontal configuration emplacement drifts is possible in a reason-able period of time.

Jan 1, 1988

By Changwoo Lee, Jan M. Mutmansky

"This paper results from the research at The Pennsylvania State University into establishment of standard procedures for characterization of respirable coal dust performed by the Generic Mineral Technology Center for Respirable Dust and supported by the U.S. Bureau of Mines and the Mining and Mineral Resources Research Institute at The Pennsylvania State University. The research performed within this particular project is oriented primarily toward standard methods of sampling and characterizing respirable coal mine dust underground so that the chemical, physical, and morphological characteristics are known and can be related to the incidence of coal workers' pneumoconiosis (CWP) in Pennsylvania.The research was initiated by con¬ducting a search of the literature to determine what coal and geologic variables may be related to CWP. The effects of coal rank, free silica content, the mass of respirable dust, and the level of trace elements were the major variables studied. It seems to be evident that the free silica content, the rank of coal, and the mass of respirable dust are the major concerns in defining the causes of CWP. However, based on the literature survey, the authors have concluded that verification of the causes of CWP is a multivariate problem and must be analyzed in an appropriate statistical manner.The results of a multivariate analysis of ten trace elements recorded on Pennsylvania coals are presented. Five of these trace elements were found to be of use in differentiating one coalfield from the others. Of these five elements, chromium and zirconium showed a strong positive correlation with the rank of coal while strontium showed a negative correlation.The design of the field sampling work and its application in future research is the final topic discussed. The sampling plan is oriented toward continuous miner sections with eight-stage impactors as the collection equipment. All of the field work will be conducted in the coal measures of the Allegheny Formation and above with both dust and channel samples being collected in the same mines.Coal workers' pneumoconiosis has been an important topic of concern to miners, mining engineers, and epidemiologists for decades. The general cause of the dis¬ease, the inhalation of coal mine dust, is well known, but the specific coal and geologic variables that affect the incidence and severity of the disease are still subject to debate and study. The primary objective of this study is to identify those variables of Pennsylvania coals that contribute to the CWP problem. To accomplish this task, a field sampling program is to be conducted to relate the characteristics of the coal dust, the coal seam properties, and the geologic variables to the occurrence of CWP.The first year of this study has been involved with both a literature review and experimental work oriented toward identification of causal variables related to CWP. The results of this work are presented here along with plans for implementation of the findings in future research.The previous studies of CWP causal variables have been conducted by scientists, engineers, and medical personnel. Many different types of studies have been performed with emphasis on interpreting the effects of dust mass concentration, mineralogy of the coal, the rank of the coal, and trace element concentrations upon the incidence of CWP in various coalfields. Most of the research has been conducted during the last two decades, with the British Isles, West Germany, and the United States being the primary countries involved.The review of the literature was conducted to determine which variables are likely to warrant further study in Pennsylvania. Previous research studies have related CWP to free silica and other mineral content, the rank of coal, the mass of dust in the working place, and the trace elements in the coal and lung dust. In the following sections, each of these topics will be discussed and analyzed."

Jan 1, 1988

By T. S. Mohibatsela, R. V. Ramani, C. V. Manula

The Mineral Engineering Department of The Pennsylvania State University has developed a suite of computer models for anthracite open pit mining (Fig. 1)1 Individual sub-systems have been developed to (a) interpolate geologic field data to obtain coal tonnages, waste volumes, and strip ratios; (b) simulate equipment performance on selected pit profiles to generate data required to choose coal and waste removal equipment; and (c) evaluate selected cost factors to determine either a rate of return for a specified selling price or vice-versa. This paper describes the application of two of these subsystems, namely, the Reserves Model (GEOL) and the Open Pit Materials Handling Simulator (OPMHS), and how these may be applied for the selection of equipment for a prime anthracite mining site.2

Jan 1, 1981

Researchers at National Institute for Occupational Safety and Health (NIOSH) are investigating how fiber-reinforced shotcrete mix designs can be extrapolated from proven designs employing wire mesh. This paper explains the results of round determinate panel tests and relates the flexural strength and toughness of fiber-reinforced shotcrete to known energy absorbing properties of wire mesh. The energy absorbed with central deflection is strongly influenced by the amount of fibres. The higher the energy absorption the more load cracked shotcrete can support. Two types of fiber-reinforced shotcrete were investigated, one with steel fiber and another with polyfiber. The residual load supported in Round Determinate Panel Tests can be matched with the strength of wire mesh, thus providing equivalent bag strength values for fiber reinforced shotcrete panels. There is also evidence of a panel crack width comparison to the load holding capabilities of the fiber in the shotcrete mixes. In practice, this provides designers with a method for comparing the ground support provided by fiber-reinforced shotcrete with that supplied by wire mesh.

Jan 1, 2010

By C. Clark, R. Pakalnis, B. Seymour, L. Martin, M. Stepan, M. Roworth, C. Caceres

The National Institute for Occupational Safety and Health (NIOSH), Office of Mine Safety and Health, Ground Control Engineering Branch is investigating the use of shotcrete in weak rock mass mines with the objective of reducing fatalities and injuries resulting from rock fall accidents. When shotcrete is used as part of a multielement ground support system, there is a need to know the support characteristics that the shotcrete contributes to the overall system. To quantify the support provided by the shotcrete, flexural strength tests were conducted with two common, commercially available fiber-reinforced shotcrete (FRS) mixes using the round determinant panel (RDP) test method, ASTM C 1550. A portable RDP test machine developed by NIOSH researchers was used to determine the peak flexural load and residual load capacity for a steel FRS and a synthetic poly FRS that were sprayed using dry mix procedures and equipment. Besides the flexural strength and loading behavior determined from these tests, a method was also developed for measuring the width of cracks exposed on the underside of a shotcrete panel during a RDP test and relating these measurements to residual load values. Quantifying the peak flexural load and residual loads of a shotcrete mix through on site testing at a mine and visually assessing the loading cycle and load-carrying capability of the shotcrete applied to underground workings will improve mine safety by providing a better assessment of the stability of shotcrete supported entries.

Jan 1, 2010

By Ronaldo Luna, Donald Deardorff

"The use of load and resistance factor design (LRFD) is well documented since the 1960s for structural applications when it was first adopted into the American Concrete Institute Building Code in 1963 and was established for geotechnical applications by the Danish Geotechnical Institute about the same time under the name of Limit State Design (LSD).The Federal Highway Administration required all federally funded projects to use the LRFD method by 2007. Due to this funding requirement, recent LRFD research for geotechnical applications has focused on common bridge foundations (i.e., driven or drilled shafts), resulting in good documentation for design using these methods. Unfortunately, the practicing geotechnical engineer does not intuitively know how to create resistance factors for deep foundations systems that have design methodologies and installation techniques significantly different from driven or drilled shaft foundation systems.Helical anchors typically use semiempirical methods for preliminary capacity design and a torque correlation methodology during installation to verify required capacity. Here we provide basic information on the LRFD methodology for helical anchor design. Additionally, the First- Order Second Moment (FOSM) method is applied to a previous helical anchor study (Hoyt and Clemence 1989) to determine resistance factors for an ultimate tension capacity strength limit state.MethodologyThe concept of the LRFD method is to establish factors for both the load and resistance that account for uncertainties using probability theory. The factors increase the nominal load and/or decrease the nominal resistance such that the probability of failure is an acceptable value. LRFD also accounts for the effects of load combinations (dead, live, seismic, wind, etc.) by determining the probability of the combined occurrence. The general form of the LRFD equation used by American Association of State Highway and Transportation Officials (AASHTO) is:"

Jan 1, 2009

By L. M. English

In the past 30 years, the research arena has changed significantly in all fields, mining included. The relative representation of the topics of research, how the research is performed, and who has performed the research have all changed significantly over these decades. Although there are a lot of subjective ideas concerning these changes, and there is a fair amount of anecdotal "analysis," there has been no actual study of these trends to verify the facts. The period studied is divided into six five-year periods for the purposes of comparison. Some of the trends analyzed are as follows: ? Co-authorship ? Research areas ? Research sponsors ? Cooperative efforts

Jan 1, 1995

By Joseph D. D?Annunzio

The report on cut and cover construction as assembled by the Organization for Economic Co-operation and Development covered 17 different countries based on a total of just under 400 individual replies of which about 30% of the replies were from the United States. In general, the review of replies to the questionnaire led to the impression that cut and cover construction is a very inadequate method which needed many improvements. This unfavorable judgment, however, resulted in the manner in which questions were asked rather than in the method itself, It can be concluded that the cut and cover construction method has proved itself as an excellent method and that it has reached a high degree of technological perfection, but inadequacies stilt exist, and for purposes of complete evaluation, comparisons with mining tunnelling methods should be carried out. The principal problem in cut and cover construction is the interaction during construction between the construction process and the existing use of the surface and subsurface environment. The greatest impediments to cut and cover construction have little to do with the physical works being constructed, but more with creating a site in which the work can be performed. Utility relocation, traffice maintenance, and protection of existing structures are major problems for the construction contractor, and too frequently receive insufficient attention in preconstruction planning and design. Some OF the areas in which research should be undertaken are as follows:

Jan 1, 1970

By Eugene F. Casey

An immersed tube tunnel is one which is composed of elements that are prefabricated in a dry dock or on shipways, usually away from the actual tunnel site, floated into position and then lowered into a preexcavated and graded trench below a body of water. The elements or segments are then connected to each other under water, and the entire immersed tube tunnel covered with a protective blanket. Because water is essential, this method of construction offers advantages in a relatively limited field. By the very nature of the method, lowering elements to the bottom of an underwater trench, immersed tube tunnels are primarily used for crossing bodies of water, such as rivers, bays, straits, canals, etc. Although this method of construction was first used for a subaqueous railway crossing under the Detroit River in 1906-1910 and again in 1912-1915, for a four track subway tunnel under the Harlem River in New York, the reason immersed tube tunnels construction has not been used more extensively is simply a matter of economics. Bridges, compressed air tunnels and similar structures could be constructed at a lower cost.

Jan 1, 1970

By Lon Santis

This paper analyzes 110 serious accidents that occurred during the use of commercial explosives in mining, construction, and other industries since 1992. Data from the Mine Safety and Health Administration (MSHA), the Institute of Makers of Explosives (IME) members’ reports, and media reports were used. Only accidents that resulted in a fatality, significant lost work time, or injury to the public were considered in this analysis. The accident data were analyzed for trends in frequency, type of industry, cause, type of individual harmed, and geographical region. Although the data showed that explosives accidents are decreasing in frequency, still, 29 people were killed and 134 injured. Eightyfive of the accidents occurred in mining, 17 occurred in construction, and 8 occurred in other industries including oil recovery, seismic exploration, agriculture, logging, boiler cleaning, and avalanche blasting. As other studies have shown, flyrock and failure of blast area security caused over half of the accidents. Mishandling, misfires, or fumes caused just over one-third of the accidents. Lightning and drilling into explosives were responsible for the remainder of accidents.

Jan 1, 2003

By J. Mutmansky, C. Lee

"The size distribution of coal mine dust is one of its most important characteristics because the particle size is known to govern the region of deposition of dust within the human respiratory tract. In addition, dust characteristics, particle settling rate, transportability of the dust and the efficiency of dust control measures are all affected by the particle size. As a result, particle size distribution of the airborne dust in a coal mine is a fundamental property that impacts on the health of underground workers.This paper deals with some basic findings on the size distribution of cool mine dust based upon the results of data collection in six underground coal mines using continuous miners. All airborne dust sampling was performed on development sections using eight stage cascade impactors and permissible 2. O l / min pumps. Samplers locations were fixed in the sections with respect to the continuous miner and roofbolter or with respect to the intake and return airways • Sampling times varied from 45 minutes to 6 hours , depending upon the dust concentration. The first topic of analysis was the test to determine whether the commonly used lognonnal distribution is valid when applied to airborne coal mine dust. Secondly, the conventional method of representing the· distribution using two parameters, the geometric mean and the geometric standard deviation, is questioned. Since almost all size data were found to show some degree of bimodality, the two parameter lognonnal distribution may not represent the total airborne coal mine dust size , quite adequately. Bimodal lognonnal distribution was tested to be a good approximation in the case with two modes in the size distribution.The final topic discussed in this paper is the 'locational variation of the airborne dust size distributions within the continuous mining section. The size distributions varied considerably at the different sampling locations with the continuous miner and roofbolter producing the coarser distributions. Fine size modes in the range of 2 to 4µ m including mostly respirable dust particles were present in almost all samples • Second dust generating sources such as the shuttle car movement were found to be major causes of high respirable dust concentrations in some cases ."

Jan 1, 1988

By J Janus, J Krawczyk

Local accumulation of methane in workings of a longwall is one of the most common causes of mining accidents. This gas has many sources: it may leak from goaf through permeable sealings, or through gaps between roof support sections. This stream is then mixed with ventilation air. In the case of auxiliary ventilation ducts, or free jet fans, we have to consider the propagation of a clean air stream in areas with an air-methane mixture. Therefore, modelling the mixing process of methane with ventilation air is important for determining the hazardous areas and thus to ensure safety. Computational fluid dynamics (CFD), specifically the finite volume method, provides tools for such modelling. However, this method requires validation. For this purpose, we have compared measurements with simulations for selected test cases having sufficient similarity to the present problems. In typical methane explosion hazard areas, such as a longwall, it is difficult to make measurements sufficiently to confirm credibility of 3D CFD simulations. As a relevant alternative, an injector methane drainage station placed in an underground gallery has been selected. The stream of drained goaf methane is entrained by compressed air powered injectors and then mixed with ventilation air. Research in this area does not interfere with mining operations, it can be performed safely in a repeatable way. Flow within the site has many similarities to conditions in mine drifts of the longwall area. Comparative research of measuring the methane concentration distribution in the mine drift downstream injector station, with simulations of propagation of the methane stream, was carried out. For measuring the methane concentration distribution in the mine drift a multipoint measurement velocity system developed at the Strata Mechanics Research Institute was used. For the simulation a finite volume method for turbulent flow of gas mixtures was applied. Preliminary results of comparison are shown as graphs and tables.CITATION:Janus, J and Krawczyk, J, 2013. An analysis of the mixing of air and methane in the stream produced by the mine injector station - Presenting preliminary results of measurements and modelling, in Proceedings The Australian Mine Ventilation Conference , pp 49-56 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 1, 2013

By Douglas Rudenko

Have you ever had a neighbor complain about a blast one day, but says the next day’s blast was better, even though the Peak Particle Velocity (PPV) increased? How about neighbors that complain about a small trimming shot, but don’t complain about a big production shot? Have you ever reduced the pounds/delay and gotten more complaints ? Mine operators are often bewildered when trying to correlate blast design parameters to homeowner complaints. When complaints come, standard practice is to set up a seismograph in the neighbor’s yard to monitor the blast. Immediately after the blast, the neighbor usually reappears from the house wanting to know the seismograph reading. When told that the PPV was 0.1 in/set, well below damage levels, the homeowner explodes saying “Your machine is wrong”. Perhaps he is right. Understanding why people complain about blasting is not about PPV, it’s all about structural response. People could care less about how much the ground vibrates; they only care about how much their house vibrates. In this paper we will examine the primary factors that control structural response. Through the use of a large array of digital seismographs, the effects of two production blasts on a community surrounding a quarry are studied. The predicted structural response resulting from each blast is calculated at 175 seismograph locations. The predicted response is then correlated to previous research on human perceptibility to vibrations. The results of this analysis and correlation give a better understanding of what the residents actually feel in their homes. The detonations of two single-hole signature blasts are also recorded by the array of seismographs revealing the dynamic characteristics of the geology ‘surrounding the quarry. Once the influence of the geology was understood, blasts were configured to disrupt the natural rhythm of the site through the use of millisecond delays.

Jan 1, 2000

By Hemant Agrawal, Arvind Kumar Mishra

Rock breakage using blasting is one of the important operations of the mining industry. The delay timing decides the initiation sequence of holes in the blasts and is a crucial factor to improve the overall blast performance. Blast-induced ground vibration is an undesired yet unavoidable outcome of blasting. Peak particle velocity (PPV) is commonly used to measure the magnitude of blast-induced ground vibration. Channelization and better utilization of explosive energy delay are provided between blast holes using initiating systems. The use of different initiation systems even on the same site results in a very significant difference in the values of PPV generated. Due to scattered delay timings of pyrotechnic detonators, the scaled distance predictor equations obtained become unreliable. The error/scatter in delay timing leads to overlapping of holes and hence increases in maximum charge per delay. In this paper, an analytical approach has been discussed to measure the overlapping of holes during blasting with initiation systems having inherent cap scatter. The analytical approach will be calculating the probable overlapping of holes due to scatter in delay timing and will help blasting engineers to understand the variation in maximum charge per delay due to pyrotechnic based initiating systems. In this study, a total number of 102 blasts were conducted using different initiation systems at two different opencast coal mine and 221 corresponding blast vibrations events were recorded.

Nov 10, 2020

By G. J. Schneider

The higher stress zone adjacent to an underground opening causes a velocity anomaly that can be mapped in favorable conditions with an acoustic cross-borehole logging system. Bureau personnel used a specialized system at the Colorado School of Mines test site near Idaho Springs, Colorado, to locate the Edgar Mine tunnel and its associated stress anomaly. Data from the logging system' were processed with a tomographic image reconstruction algorithm to map the opening and the location of the high velocity zone around the opening. The velocity imagery from tomographic analysis compared favorably with predictions from numerical modelling of the location of the stress anomaly.

Jan 1, 1988

By C. Allen, H. Zhang, L. Falk

Shock loss occurs when there is a change in the area of a drift or airflow direction. These shock losses accumulate throughout the ventilation system and will increase the mine resistance as underground development deepens or becomes more complex. Hence, shock losses should be carefully considered when establishing the resistance through a series of branches as they can be a significant contributor to the estimated mine resistance. Using computer software to simulate a ventilation system is common but applying reasonable shock loss factors in the software is not a simple or direct process. Determining shock loss values includes the consideration of historical measurements, application of factors and formulas from literature based on the geometry and complexity of junctions, and the geometry variation at different locations. For example, the shock loss at the bottom of a ventilation raise can consist of a 90° bend and a sudden area change. This paper presents an approach for estimating shock loss factors near raise junctions using computational fluid dynamics (CFD). Current and future work will also be compared with field measurements. Results from this approach are compared with software defaults as well as classical techniques in published literature and standards.

Sep 1, 2024