Search Documents

By J. A. Organiscak

Two types of airflow splitting methods for improving longwall dust control were investigated by NIOSH's Pittsburgh Research Laboratory. These methods included a translucent mesh barrier and a staged spray barrier system to confine the shearer-generated dust to the coal extraction side of the long wall face while maintaining a cleaner split of airflow on the opposing worker walkway side of the longwall face. The translucent mesh barrier was shown to provide notable dust control effectiveness up to 60 m downstream of the shearer in both the laboratory experiments and a long wall field evaluation, provided the mesh barrier remains relatively parallel to the airflow. The staged spray barrier system showed marginal dust control effectiveness in the laboratory and subsequently was not field tested.

Jan 1, 1999

By Gary E. McIntosh

This paper will discuss a U.S. Bureau of Mines project designed to assess and predict the impacts which surface coal mining will have on the regional shallow groundwater systems in the Powder River Basin. The primary objective will be met through the accomplishment of the following: (1) The design and establishment of a regional network for monitoring of shallow groundwater systems in the Basin. (2) Monitoring of the direction and rate of flow, water quality, and aquifer characteristics of the shallow groundwater systems. (3) Development of a regional groundwater simulation model with appurtenant subroutines to determine local effects.

Jan 1, 1978

By W. E. Lytle

Environmental permitting is a critical component for joint-venture projects located in the Commonwealth of Independent States (CIS). This paper presents the basic CIS regulatory environmental requirements for permitting of a hypothetical mining source and emphasizes air permitting. Topics which will be covered include: the Expertiza process, air pollution quantification, data requirements, and air pollution modeling protocol. A comparison will be made between permitting requirements for air pollution sources in the United States and a CIS country, Kazakstan.

Jan 1, 1998

By C. Newman

"Over the past decade, as conventional underground deposits have been depleted, mining operations have been forced to produce at greater depths and in more geologically and geometrically challenging conditions. As such, there has been a global increase in the application of cemented paste backfill (CPB) in tabular deposits utilizing open stope mining method with a delayed backfill placement. Despite the extensive use of CPB, many fundamental factors affecting the design of safe and economical fill structures are still not well understood. A critical issue in the design of backfilled stopes is the determination of stress states within the fill material itself as well as the surrounding rock mass. Analytical equations provide a means of quickly evaluating the effectiveness of a given design. However, in developing these equations large assumptions are implemented to simplify the design problem. It is important that one understands these simplifications as well as their effect on the overall design. This paper investigates common analytical equations utilized in the evaluation of single vertical backfilled stopes and their assumptions through comparisons to numerical modeling results. INTRODUCTION AND BACKGROUND As the mining industry continues to produce at greater depths and in more geometrically and geologically complex conditions, cemented paste backfill (CPB) has gained traction as a means of providing localized ground support in modern stope-and-fill mining operations. The application of CPB in stope-and-fill mining as a ground support material has led to the reduction of mine waste costs, while increasing both mine production and stability. In an attempt to achieve total extraction of the mining reserve, stope-and-fill mining operations often employ a primary-secondary excavation-support sequence (see Figure 1). In this practice, primary stopes are initially excavated and then backfilled in a two-stage pour; plug and final pours. During the first stage (plug), a CPB material with increased cementitious content is placed at the bottom of the stope providing a strong base from which to build a backfill structure as well as protect mine works from material breakthrough. Following placement, the backfill plug is allowed to cure to a given strength as designated by the mine design. Upon achieving the required cure strength, a less cementitious CPB material is used to backfill the remaining stope area. This is referred to as the final pour stage. With all primary stopes excavated and supported using backfill, the extraction of the secondary stopes commences as primary stopes are utilized as artificial pillars. While the application of CPB in a primary-secondary extraction-support mining sequence allows for a substantial increase in reserve recovery, it is imperative the mining personnel and planning engineers have a clear understanding of the material’s behavior and stress interaction with the surrounding rock mass to ensure the most efficient, cost-effective, and safe extraction of underground deposits."

Jan 1, 2018

By Ihsan Berk Tulu, Ted Klemetti, Deniz Tuncay

"The Analysis of Retreat Mining Pillar Stability (ARMPS) program was developed by the National Institute for Occupational Safety and Health (NIOSH) to help the U.S. coal mining industry to design safe retreat room and pillar panels. ARMPS calculates the magnitude of the in-situ and mining-induced loads by using geometrical computations and empirical rules. In particular, the program uses the “abutment angle” concept in calculating the magnitude of the abutment load on pillars adjacent to a gob.In this paper, stress measurements from U.S. and Australian mines with different overburden geologies with varying hard rock percentages were back analyzed. The results of the analyses indicated that for depths less than 650 ft, the ARMPS empirical derivation of a 21° abutment angle was supported by the case histories; however, at depths greater than 650 ft, the abutment angle was found to be significantly less than 21°. In this paper, a new equation employing the panel width to overburden depth ratio is constructed for the calculation of accurate abutment angles for deeper mining cases.The new abutment angle equation was tested using both ARMPS2010 and LaModel for the entire case history database of ARMPS2010. The new abutment angle equation to estimate the magnitude of the mining-induced loads used together with the LaModel program was found to give good classification accuracies compared to ARMPS2010 for deep cover cases.INTRODUCTIONIn the early 1990s, the Analysis of Longwall Pillar Stability (ALPS) was introduced by Mark (1990) as a chain pillar design software and was generally accepted and used by the U.S. coal mining industry. Following the success of ALPS, the National Institute for Occupational Safety and Health (NIOSH) developed the Analysis of Retreat Mining Pillar Stability (ARMPS) program for designing retreat mining pillars using a similar approach as ALPS (Mark and Chase, 1997). The Australian mining industry also recognized the success of ALPS, and Colwell, Frith, and Mark (1999) calibrated the program to Australian conditions. The ALPS and ARMPS programs draw their strengths from the large databases that are used to calibrate them (Mark, 2009). However, following the Crandall Canyon Mine collapse in 2007, NIOSH had to reconsider the pillar design criteria used in deep-cover retreat mining (Mark, 2010). The ARMPS overburden load prediction algorithm was improved to more accurately predict the loading of narrow panels with high overburden depths by implementing the pressure arch concept, and this new version is called ARMPS2010."

Jan 1, 2019

By O. H. Rostad

The importance of geology in ore reserve estimation has been stated by several authors. King et, a1 (1982) reported that the geological factor was the key deficiency where serious weaknesses in ore reserve estimation appeared, especially in grade. The foundation for a successful mining operation is an ore reserve estimate that neither overstates nor understates the true situation as to tonnage and grade. Overestimation of reserves can have devastating economic results. Overestimation of tonnage may lead to wasting money on an overbuilt plant or worse yet, if grade is seriously overestimated also, to undertaking an operation that is doomed to economic failure from the outset. Understated ore reserves can result in building a plant that is smaller than optimum which may result in higher unit costs. At worst, understated tonnage of ore reserves can result in a missed opportunity for a profitable operation. Failure of the ore reserve estimator to properly take into account all geologic and allied factors pertaining to the deposit can result in an ore reserve estimate that is overstated either as to tonnage or grade or both. This paper will focus on these factors and will not discuss rules about distances above or below a drift that should be given a classification of proven or probable, etc. Each deposit has its own particular characteristics which often require modification of general rules for defining limits of proven and probable ore blocks that work satisfactorily in another deposit. A number of references are included which discuss some of these problems. Definitions of ore reserve classifications and an illustration are given in the Appendix.

Jan 1, 1985

By J. M. Keane

SMEThis loaner is concerned with important criteria relative to the design of small flota¬tion concentrators. For the purpose of this discussion a small flotation concentrator is defined as a facility having a designed ore feed rate of 1,000 metric tons per day (MTPD), or less. The decision to design and construct a small flotation concentrator is the culmination of a series of logical steps relative to the ex¬ploitation of a potential ore body. The information required for a rational and timely decision regarding plant design and construction is usually contained in a feasibility report. The information required in a thorough feasibility report has been summarized (1) (2) as follows: Type of mine or plant and proposed daily, monthly, and yearly production of all intermediate and final products. Quantity of indicated ore reserves. Location of project illustrated by a map showing surroundings, trans-portation facilities, and existing utilities.

Jan 1, 1980

By H. Soto, J. Jara, F. Nava, J. B. Hiskey, B. J. Scheiner

Laboratory tests showed that ozone can destroy free cyanide and weakly complexed cyanide (copper and zinc cyanides). The iron cyanide complexes were resilient to ozonation but they were effectively eliminated by precipitation with Fe (II) ions under controlled pH conditions. Tests carried out with effluents from a number of gold mills confirmed the results obtained previously in the laboratory. The industrial samples contained cyanide in the form of free cyanide and also complex copper, zinc and iron cyanides. Solutions that contained up to 184 mg/L total cyanide, after the ozone treatment analyzed only 2-3 mg/L dissociable cyanide and 50-90 mg/L CN(t) (accounted for by iron complexes). After precipitation with ferrous ions total cyanide was reduced below 1.5 mg/L. For most solutions ozone consumption ranges from 3 to 6 g per g of cyanide destroyed. However, certain impurities (such as SCN and S203 ), present in some effluents, increase ozone consumption and make it more difficult to reduce copper concentration to satisfactory levels.

Jan 1, 1993

By A. M. Esteves

This paper addresses the question ?What constitutes the right level of investment in community goodwill?? Mining companies are still grappling with this subject even though the business case for community investment has been accepted at a general level, as a strategy for competitive differentiation, risk mitigation and a desire to deliver - and to be seen to deliver - a ?net benefit? to affected communities. Gaining senior management commitment to long term social projects, by definition characterized by uncertainty and complexity, is made easier if projects are shown to benefit the site?s strategic goals. Case study research amongst mining companies, including interviews with social investment decision-makers, has assisted in developing a SI-DAT (Social Investment Decision Analysis Tool), a decision model for evaluating social projects in order to create value for both the company and community. Multi-criteria decision analysis techniques integrating business planning processes with social impact assessment have proved useful in assisting mining companies think beyond seeking access to required lands and peaceful relations, to how they can meet their business goals and contribute to regional development.

Jan 1, 2007

By Osamu Shoji, Mamoru Kato

In order to continue improving the quality of telecommunication services to meet the expectations of the people and to increase reliability in the case of emergency such as disasters, Nippon Telegraph and Telephone Public Corporation (NTT) is now constructing cable tunnels which connect main urban telephone offices. However, almost all the large cities in Japan are built in river deltas, so that their chief underground features are low strength fluvial deposit and a high ground water level. In this saturated soft ground, the ground surface at an installation site has occasionally subsided through deep excavations for cable tunneling construction. As these settlements have sometimes caused accidents, especially in densely populated large cities, NTT has made much effort to avoid such accidents. These accidents have occured mainly in flexible sheeting due to its inward movement. Therefore, after the investigation into forecast method of sheeting deflection with accuracy, NTT has reserched a system which forecasts ground surface settlements. This paper describes an outline of the ground settlement forecast system and practical example on the site.

Jan 1, 1979

By T. S. Golosinski

Use of contractors to excavate, load and haul mined material is widespread in Australia. It is also becoming a common practice in many other countries. The paper reviews current practices in Australian contract mining. Various types of contracts are reviewed, including partnering arrangements and alliancing between the contractor and the mine owner. Applicability of various types of contracts to specific mining situations is discussed, as is their impact on overall mine performance

Jan 1, 1998

By Hun-Joon Sohn

Chemical metathesis and electrochemical reductive conversion of chalcopyrite in the presence of cupric ions were studied in sulfuric acid solutions for temperatures up to 90°C. The process results in the displacement of iron in the chalcopyrite lattice by copper ions with rejection of ferrous ions to the aqueous phase. The rate of reaction was found to be very slow, requiring subsieve-sized particles for extensive reaction to occur. The reaction was insensitive to copper ion concentration. Stirred ball milling of chalcopyrite introduces lattice strain. The reaction rate with strained particles was markedly higher than with annealed chalcopyrite. The reaction rate was found to be diffusion-limited, showing two distinct reaction stages corresponding to the formation of CuS (covellite) followed by Cu1.8S (digenite). The experimental activation energy for the first stage was 102 J (24.4 kcal) and the second stage 123 3 (29.4 -kcal) per mole. The rate-controlling process is diffusion of lattice ions, probably iron. The enhanced rate after milling is attributed to a recrystallization process which provides new paths for diffusion along dislocations and grain boundaries.

Jan 1, 1984

By Walter C. Jr. Sherman

6.1-SOIL MECHANICS WALTER C. SHERMAN, JR. 6.1.1-DEFINITION Soil mechanics is concerned with the behavior of soils in relation to the design, construction and performance of engineering works. In this context, soils include all uncemented materials, such as, gravels, sands, silts and clays, either naturally or artificially deposited, and, under certain conditions, weakly cemented, badly broken or weathered rocks. Many types of mine wastes have physical properties similar to those of soils and thus are amenable to analysis by the principles of soil mechanics. The principles of soil mechanics are used to solve such problems as the stability of various types of structures with respect to failure of the supporting soils, determination of soil pressures and deformations in the design of foundations and retaining structures, analyses of seepage and under seepage, and stability with respect to sliding of excavated and natural slopes and embankments. 6.1.2--SOIL PROPERTIES Soils are classified on the basis of index properties, such as, particle-size distribution and plasticity characteristics. The Unified Soil Classification System1 used to classify soils for engineering purposes is shown in Table 6-1. Other physical properties important from an engineering standpoint include natural water content, density, permeability, shear strength and compressibility. Physical properties are evaluated on the basis of carefully conducted laboratory tests. Procedures for laboratory testing are presented by Lambe. 2 Soils, either of natural or artificial deposition, are complex three-phase materials consisting of solids, water and gas (usually air or water vapor). One of the most important principles in soil mechanics is based on the concept of effective stress in the soil-water system. A stress, a, applied to a saturated soil is taken partly by the soil particles (effective stress, a'), and partly by the water in the voids (neutral stress or pore water pressure, u,) so that a' = a - u,. Only the effective stress affects the mobilization of frictional resistance between soil particles. In partially saturated soils, the pore-water pressure is negative due to capillary effects. The effective stress relation remains valid for small negative pore pressures, but anomolies arise if the soil is significantly unsaturated. In this case, the pore pressure is commonly assumed as equal to the pore air pressure, u,, so that [ u' = u - u..] 6.1.3--SOIL EXPLORATIONS Soil explorations are carried out by borings or test pits. Hvorslev3 has reviewed procedures and boring requirements, such as, type, depth and spacing employed. In general, soil strata are logged and representative exploratory samples are recovered for classification purposes and for determining the subsurface soil profile, which is constructed utilizing also knowledge of the geology of a site. Undisturbed samples are recovered if the properties of the undisturbed soil are required, as generally is the case. The undisturbed soil sampler consists of a thin-wall tube

Jan 1, 1973

By T. J. Batchler

"INTRODUCTION A variety of standing supports are being applied in underground coal mining operations to assist in providing stable mine entries. Standing support performance requirements depend on the nature of the ground deformation behavior, and therefore require a good understanding of the interaction between the support and ground to achieve an optimum roof support system. Mining operations often employ standing support strategies that are unique to their mine. A support system that functions successfully in one situation could fail in another situation. It is important that the design and installation of the standing support systems provide the highest probability of preventing roof falls. Considerable research and development has been conducted by support manufacturers to develop standing support technologies for the mining industry. The National Institute for Occupational Safety and Health (NIOSH) researchers, with the cooperation of various support manufacturers, have developed a testing protocol that evaluates performance characteristics of these standing support systems.. Using this protocol, each of these support systems has gone through rigorous safety performance testing under controlled loading conditions in NIOSH’s Mine Roof Simulator (MRS). It is important to understand the unique performance characteristics of each type of standing support. This information allows mine operators to determine if the support system is compatible with the conditions in the mine and to compare each support system and determine if that support will provide equivalent or improved support capability. This paper provides a comprehensive summary of support performance to facilitate this objective. SUPPORT DESIGN CRITERIA AND THE GROUND REACTION CURVE Standing roof support systems are designed to prevent unplanned roof falls. The selection of the standing support is based on the interaction of the support with the surrounding rock mass, and its function is to assist the primary support system in maintaining the integrity of the immediate roof. The standing support works to control the immediate roof structure by resisting deformation that contribute to delamination of the bedded layers and displacements that lead to separation of the bolted zone from stable roof structures above the bolted horizon (Barczak, 2005). Standing supports should be installed early enough and develop sufficient capacity to bring deformation of the ground into equilibrium before a critical deformation is reached that leads to a detached block condition, at which point the standing support will be required to carry the entire dead weight of the detached rock mass to prevent a roof fall from occurring (Mucho, 1999). The ground reaction curve (GRC) can be used to assist in determining the compatibility and design requirements for a support system. It is important to understand the significance of the ground reaction relative to the support design requirements. The ultimate objective is to match the support behavior to the ground response and create equilibrium (stability) of the rock mass. This equilibrium is accomplished when the support loading intersects the ground reaction curve (Figure 1). Essentially, the ground reaction concept implies that the convergence in the mine entry is controlled by the magnitude of the support resistance. However, as a passive support, the standing support develops its loading through convergence. As such, the support needs to maintain its load capacity through the associated deformation of the ground and resulting entry convergence until the equilibrium condition is reached. This requirement causes factors such as the support stiffness, peak capacity, yielding characteristics, and stability to be fundamental to proper support design and application. NIOSH contributes to providing this understanding through full-scale performance testing of support products in the Mine Roof Simulator."

Jan 1, 2017

By Philomene Verlaan

The Code for Environmental Management of Marine Mining (the Code) was adopted in 2001 by the International Marine Minerals Society (IMMS) following extensive and wide-ranging consultation, and updated in 2002. The Code provides a framework and benchmarks for development and implementation of an environmental program for a marine exploration or extraction site by marine mining companies, and for stakeholders in governments, NGOs and communities in evaluating actual and proposed applications of environmental programs at marine mining sites. The Code provides for its review by IMMS after five years, with a focus toward possible modifications based on experience with its implementation and in light of developments in the field of marine mining and associated environmental practices, in consultation with the marine mining industry and with other stakeholders at marine mining operations. As the review of the Code is now due, to begin the process, this paper will present an analysis of the Code and initial suggestions for its modification. Keywords: marine mining, marine environment, sustainable development, marine environmental risk management

Jan 1, 2008

By K. H. McDaniel

An interactive mine ventilation simulation software program (WIPPVENT) was developed at the Waste Isolation Pilot Plant (WIPP). The WIPP is a U.S. Department of Energy (DOE) research and development project located near Carlsbad, New Mexico. The facility is designed to provide a permanent, safe underground disposal of U.S. defense generated transuranic waste in bedded salt. In addition to its regular functions, the underground ventilation system is engineered to prevent the uncontrolled spread of radioactive materials in the unlikely event of a release. To enhance the operability system, Westinghouse Electric Corporation has developed an interactive mine ventilation simulation software program (WIPPVENT). While WIPPVENT includes most of the functions of the commercially available simulation program VNETPC (copyright © 1991 Mine Ventilation Services, Inc.), the user interface has been completely rewritten as a Windows ® application and screen graphics have been added. WIPPVENT is designed to interact with the WIPP ventilation monitoring systems through the site wide Central Monitoring System.

Jan 1, 1996

By S. L. Bessinger

Introduction In the design and specification of a ventilation system for an underground mine, it is necessary to make reasonably accurate estimates of the pressure losses in the various airways of the mine. These estimates can be made with little difficulty for open airways with simple geometric cross-sections, such as those cut by continuous miners or tunnel-boring machines. The situation is much different on a longwall face, where the airway's complex geometric cross section and the presence in the airway of obstructing equipment having a variety of shapes make it difficult, if not impossible, to estimate pressure loss using traditional methods of calculation. Head losses in mine entries are calculated using Atkinson's Equation. [22H= KPLQ (English) H= KP 3Q (SO (1) 5.2AA] where H = pressure loss, in. of H2O (Pa); K = friction factor, lbf•min2/ft4 (kg/m3); P = perimeter, ft (m); L = airway length, ft (m); Q = airflow quantity, ft3/min (m3/sec); and A = flow cross-sectional area, ft2 (m2) In this equation, the friction factor, K, is an empirical constant that describes the aerodynamic roughness of the airway. Typically, the K-factor for a given airway is determined by measuring the factors H, P, L, Q and A in Equation (1) and calculating K. Tables of friction factors calculated in this way are found in textbooks and handbooks that deal with mine ventilation analysis. Unfortunately, very few K-factors have been measured on longwall faces, and the accuracies of those that have been measured are entirely site specific, because of the wide variety of equipment found on longwalls. The development of a technique for prediction without mine-site measurements of the friction factor for any longwall face, regardless of its configuration, will thus be very useful in the design of ventilation systems for mines in which longwall mining is practiced. Calculation of pressure losses using Atkinson's Equation (1) and empirically determined K-factors provides accurate and useful approximations in cases where the airways have relatively simple cross sections. However, a careful analysis using the principles of fluid mechanics shows that such calculations are based on two assumptions that are not strictly correct when there are obstructions in the airway. The first assumption is that the air velocity distribution in the cross section, particularly around the perimeter, is uniform. This assumption results from the fact that the tabulated K-factor values found in the literature are based on field measurements with uniform conditions. Such uniformity does not exist in longwall airflows. The second common assumption is that the K-factor, and corresponding head loss, is independent of the Reynolds Number (NR) for a given flow. In fact, this assumption is not strictly correct, and is particularly erroneous in the case of irregular protuberances into the airflow, such as those found on a longwall face. The errors arising from the assumptions may be avoided by using K-factors calculated by a newly devised method, described below, which takes into account the fundamental principles of aerodynamic drag analysis. This new technique has two advantages: first, it is flexible enough to model any longwall, regardless of equipment configuration; second, it employs terminology and equations familiar to those who perform mine ventilation analysis, using K-factors, for which ventilation engineers have an intuitive understanding, rather than drag coefficients. To provide guidance for development of a longwall drag model, data were taken on two modern longwalls operating in substantially different conditions. Pressure measurements at Mine B were made with 200-foot (61-m) sections of 1/8-in. (3-mm) diameter plastic tubing, attached to a Dwyer Magnehelic gauge. Pressure drops were measured in 200-foot increments down the face, and summed to give the drop for the entire face length. This technique was found to produce small, repeated errors because of the number of segments required to span the longwall. At Mine A this problem was avoided by using a single, continuous, plastic tube for the entire face length. The psychrometric properties of the air were measured for both Mines A and B. A calibrated, standard vane-anemometer was used to measure the airflow on both faces. Finally, numerous dimensions were measured on both faces, and face profile drawings were obtained to allow detailed evaluation of the face equipment geometry. From this information, accurate evaluations of the average wetted perimeter and average area of the longwall face airways were made. Since the airflow is not confined to inside the powered supports at all points along the face, a quadratically weighted average of the airflows measured at various stations along the face was calculated: [n2Qavg =Qi Ii / It(2)i=1] where [Q, avg = average airflow for analytical purposes, ft3/min (m3/sec); Q= airflow at station i, ft3/min (m3/sec); 1= length of segment represented by Q, ft (m); h= length of longwall face, ft (m): and n= number of quantity measurement stations.] The quadratic weighting scheme was chosen because the

Jan 1, 1992

By E. O. Jacobs

A coupled solid and rock fracture mechanics model is used to analyze the blast driven fracture propagation phase, predict propagation rates, and directly measure shockwave and particle velocity. The study incorporates quasi-static pre-stress conditions for different depths below the surface. The scaled experiments include dynamic material properties and quasi-dynamic fracture mechanics which are shown to significantly affect the fracture propagation. The study was performed using three-dimensional cube-type laboratory models fabricated from Polymethylmethacrylate (PMMA). The test cubes are loaded with explosives and resulting fracture networks are studied. The dynamic evaluation of the spatial crack system is monitored using a high speed image camera. An experimental model set-up is designed to indicate the pre-stress effect on deep borehole blasting and is shown that the pre-stress conditions have a decisive influence on fracture profiles and orientation. It is found that shockwaves rapidly exceed the slower dynamic fractures and a majority of cracks occur due to detonation pressure. Experimental testing using different parameters validate that explosive fracturing can predict the outcome of fracture and crack propagation in any rock type with known material properties. This can be accomplished under great stress conditions deep below ground surface. The experimental study results can be applied to a variety of industries such as mining and quarrying, tunneling, oil and gas, and construction.

Feb 23, 2014

By Lukas Walder, Roland Arnold, Raphael Zuber, Richard Gradnik

This paper presents the construction of the ventilation shaft for the single-tube, doublelane Kramer tunnel. The project is located in the Federal State of Bavaria, Germany. This road tunnel, which has been under construction since December 2019, is the key component of the western bypass of the town of Garmisch-Partenkirchen. Once completed, it will relieve the town’s central district from enormous volumes of traffic. The contractor is the joint venture named ARGE Kramertunnel consisting of BeMo Tunnelling and Subterra. The whole project will be completed and handed over to the project owner, Staatliches Bauamt Weilheim, in 2023. The opening of the entire bypass section including the tunnel to traffic is planned for 2024. The tunnel, with a total length of 3,609 m (2.24 miles) and an estimated cost of EUR 178 million incl. VAT (US-Dollar 201 million), is designed as a single-tube, double-lane structure with one traffic lane for each direction of travel. The safety issue that exists for singletube tunnels is solved by a parallel escape tunnel passable for vehicles. For tunnel ventilation, an intermediate shaft has been designed with a depth of approximately 90 m (295 feet) and an inner diameter of 4.5 m (14.8 feet). All tunnels and ducts are excavated according to the principles of the New Austrian Tunnelling Method (NATM). The geological and hydrogeological conditions, the change in the profile (from round to square geometry), and also the shaft being located at a high elevation presented a demanding technical and logistical challenge. The site arrangement with the pit bank was located in a cut, the slopes of which had to be stabilized with shotcrete and soil nails. This excavation was carried out from the fall of 2020 to the beginning of 2021 because construction operations had to be completed outside the nesting period of protected birds.

Jun 13, 2023

By B. N. Conley, K. D. Luxbacher

"The underground coal mining industry in the United States has recently seen the occurrences of several high profile, multi-fatality events. The explosions that occurred at the Sago Mine in 2006, the Darby Mine in 2006, and the Upper Big Branch Mine in 2010 have caused a ripple in an otherwise steadily improving safety record. These events transpired in the midst of an unprecedented level of government regulations and modern safety technologies. The recent increase in fatal events in conjunction with a minimal decline of both fatal and non-fatal injuries over the past decade may signify that current safety practices have reached a level of diminishing returns. Risk management, a safety approach that has been successfully applied in various industries including mining across the world, may provide a means to surpass the safety plateau in the U.S. RISKGATE is an Australian risk management program for mines that shows great potential for application in the U.S. However, fundamental differences between the coal mining industries in Australia and in the U.S. prevent direct implementation. This paper discusses aspects of the RISKGATE body of knowledge that require some adaptation before this program may be applied to the U.S. mining industry. INTRODUCTION The underground coal mining industry in the United States has recently seen the occurrences of several high profile, multi-fatality events. These events transpired in the midst of an unprecedented level of government regulations and modern safety technologies. The recent increase in fatal events in conjunction with a minimal decline of both fatal and non-fatal injuries over the past decade may signify that current safety practices have reach a level of diminishing returns. In order to combat this stagnation in performance, a modification to the manner in which the mining industry approaches worker safety may be warranted. The implementation of a risk management based approach presents a potential method for not only surmounting the current safety plateau but also for achieving the long desired objective of zero harm. RISKGATE a risk management body of knowledge designed for the Australian mining industry that shows great potential for application in the U.S. This paper discusses aspects of the RISKGATE body of knowledge that require adaptation before implementation in the U.S. mining industry. These adaptations are based on differences in operating practices and governmental regulations. This discussion will focus on the three major U.S. risk areas of explosion prevention, ground control, and moving equipment."

Jan 1, 2015