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By Lok W. Home

HISTORICAL DEVELOPMENT At the beginning of the 1970 more than 90% of pro¬duction drilling in open-pit mining was obtained with the use of rotary drills. At the same time, the use of rotary drills underground was virtually nonexistent. Several progressive mines began questioning whether efficient rotary drilling could not be adapted to under¬ground applications. In the late 60s and early 70s at¬tempts to use rotary drilling were made by Inco, Ltd. and Meramac Mining, utilizing drill units designed as a combination of small diameter raise drills and rotary drills. These initial attempts were unsuccessful primarily due to the immobility of the equipment and low degree of dependability. At the same time, experiments con¬ducted by Inco and Meramac using large diameter in¬the-hole percussion drills produced immediate encourag¬ing results. The first serious attempt at utilizing rotary drilling underground was made by Texasgulf in 1971 at Kidd Creek mine in Canada. An experimental stope 15 m wide, 36 m long, and 30 m deep was drilled using 200-mm diam rotary drilled holes. The machine used was a Robbins' Model 11 D. The Model 11D is a rotary drill designed to drill drain and utility holes under¬ground. The results of this test revealed that large diameter blasthole stoping was feasible. The indications were that the major areas of concern-fragmentation and pillar damage-could be controlled and a new low-cost stoping method resulted. At this point Texasgulf and Robbins entered into a cost-sharing agreement to develop a rotary drilling unit specifically for this stoping method. Drilling of the trial stope required pouring a concrete pad at each drill hole plus anchoring the derrick to the pad by means of rock¬bolts. Utilization expressed in terms of drilling time to available time was less than 20%. It was obvious that the required unit had to be mobile, setup time had to be minimized, and rod handling had to be improved. The first attempt at an underground production-type rotary drill consisted of mounting the basic 11D on an air-powered Eimco 600 Series crawler. The crawler was equipped with four screw-type leveling jacks operated by an impact wrench; rod rack facilities for 60 m of 1.2-m long, 170-mm diam rods; a rod handling facility consisting of an air-lift swing boom and a pipe loader which automatically positioned the rod under the float¬ing drive box; a turnbuckle anchor system requiring in¬stallation of four rockbolts in the drill drift floor and placing turnbuckles from the rockbolts to the corners of the drill; and trailing power units containing the electrics and hydraulics required to operate the drilling unit. Operation of this unit revealed a vast improvement in drilling time but some problem areas remained. The anchoring system required secondary drilling and proved difficult to adapt to the localized irregular floor surface and the trailing power unit was difficult to maneuver and was causing unnecessary maintenance costs. Addi¬tional improvements resulted in the standard 11MD rotary blasthole drill. DESCRIPTION OF UNIT This standard model consists of the drill and the power unit integrally mounted on an air-powered crawler chassis (Fig. 1). The rod-handling and leveling¬jack systems are similar to those used on the prototype Texasgulf unit and remained relatively the same. The anchoring system has been replaced by the roof jack, initially screw type but now hydraulically actuated. The anchoring system is necessary as the drill exerts more thrust than the weight of the unit. During operation it is found that the two front jacks adequately anchor the machine. The complete unit's operational dimensions are: height, 3.2 m; width, 2.1 m; and length, 4.2 m requiring a 3.5- x 3-m drill drift. The derrick is pivot mounted and raised and lowered by a hydraulic cylinder reducing the transport height to 2.5 m. The derrick is also turnbuckle mounted to allow dip angle drilling from vertical to 0.52 rad (30°) off vertical. The drill derrick can be mounted to allow dip axis to be parallel or perpendicu¬lar to the long axis of the unit. The standard rods are 1.2 m in length and are contained in racks of 10 which can be removed for transportation. Drilling torque is produced by a 56-kW (75-hp) electric motor driving a double vane pump which in turn drives a Robbins high torque gear-type hydraulic mo¬tor. In addition to the hydraulic motor, the machine drive system has a 3.9:1 planetary gear reducer. This results in six speeds with the following minimum and maximum range: Maximum rpm 112 Torque 3725.5 J (380 kg m) Minimum rpm 10 Torque 14 706.0 J (1500 kg m) Drilling thrust is provided by two bootstrap cylin¬ders pressurized by a variable volume pump. This sys¬tem can produce from 0 to 267 231.2 N thrust in either direction. Bailing air (approximately 17 m3/min at 382.2 to 490.3 kPa) is introduced through a rotary swivel pass¬ing directly through the center of the drive system to keep efficiency and losses to a minimum. Hole annulus is also kept to a minimum by using 170-mm rod when drilling a 200-mm hole. A small amount of water is added to the air for dust suppression. The machine has to be supplied with air, water, and electrical power. It does not appear feasible to provide an integral air power supply with the rotary drill such as is done with open-pit rotary drills. Air is required and installed during the primary development of the drill drift. The machine can be adapted to drill hole diameters from 170 to 228 mm for blasthole operations. The drill string consists of the rotary bit, crossover stabilizer, roller stabilizers, and standard rods. The operator controls speed and thrust pressure from an operating console which can be mounted in a conve¬nient position. Mount Isa has made an adaption to the operating console to permit remote collaring of drill holes in an uncleared floor.

Jan 1, 1982

By E. Don Dilling, W. F. Holbrook, Lloyd R. Michels, James W. Pennington, Dwight L. Harris, Wm. F. Hergert, Cyrus L. Blogett, H. A. Doerner, Ernest A. Reige

"IntroductionA process for production of magnesium by carbothermal reduction was developed at Padentheirm, Austria, by the Austro Americanische Magnesit Corporation& an independent investigation of this and other processes was started in 1937 by the Bureau of Mines at Pullman, Wash., with the cooperation of Washington State College. 10/ A preliminary study of the Radentheirm pro¬cess, as revealed by the meager information then available, demonstrated the following two major disadvantages,The use of hydrogen for shock cooling requires an enormous volume of gas - about 1,000 cubic feet for condensing each pound of metal as a pyrophoric dust which must be completely separated from this gas. Then if the gas is to be recycled for repeated use, the CO produced by the reduction reaction must be separated from the hydrogen. These separations are costly from the standpoint of equipment and operation. 11/The condensate is an impure, pyrophoric metal powder that must be purified and converted to coherent form. This is accomplished by distillation or sublimation of the metal. Distillation on a large scale involves difficulties inherent in the high boiling point (1,107° C.) and the chemical activity of the metal. To avoid these difficulties, the metal is sublimed at 600° C. and of very low pressure. This is quite easily carried out as a batch operation but is a costly procedure. A continuous sublimation process has not been accompli shed."

May 1, 1945

By A. C. Fieldner

This report, the third of a series of annual reports of research and technologic work on coal by the Technologic Branch to be released in this form, covers the fiscal year July 1, 1937, to June 30, 1938. Ever-increasing demands for information on the chemical and physical properties of coals, their proper classification rind adaptability for various uses, their economic evaluation in the complex relationships between the costs of production, distribution, and selling, and their beneficiation and amenability to conversion to more valuable and. new products of commerce have necessitated an appropriate and flexible program. To most these requirements, initiation of new projects and new physical facilities has been made.

Jan 1, 1939

By D. Harrington

For many years the Federal Bureau of Mines-in cooperation with other agencies-has been making investigations and compiling and publishing reports and statistics on accidents in the various branches of the mineral industry.4 Its assistance has been enlisted also in health, safety, and ventilation problems' incident to the driving of many large tunnels in the United States in recent years. As part of this work the Bureau has compiled considerable statistical and other information on accident experience in the tunneling branch of the heavy-construction industry. It has always been difficult to obtain up-to-date information on tunneling methods and tunnel accidents. Published accounts of tun¬nel construction furnish much interesting material of engineering, historical, or fiscal value, but usually little or no attempt is made to discuss and critically analyze the means taken to preserve the health and life of employees.

Jan 1, 1941

The success of smooth blasting essentially depends on geological condition of rock. Also, in the perimeter hole, borehole pressures with light, well-distributed explosive loads are important to minimise overbreak. In order to simulate perimeter holes with scaled model test, the borehole pressure was considered as a factor of scale. Hydraulic oil pressure is applied as borehole pressure instead of using explosives and detonators in order to remove the undesirable effects of detonators in an experimental set-up of just two perimeter holes in a tunnel blast. The borehole pressure on a perimeter hole is nearly reproduced with hydraulic pressure which is generated by non-electric detonators in the hermetical cylinder.

Dec 6, 2010

By Frank G. Schwartz

VARIABLES OF composition and environment affecting storage stability of gasoline-type fuels were studied, utilizing radioactive tracers and other analytical techniques. Changes in gasoline composition brought about by aging were determined and correlated with gum formation. The amounts of reaction of selected gasoline components, measured by radioactive tracers, were determined in a gasoline-type blend and in a mixture of 17 selected pure gasoline components. The data show that sulfur compounds, nitrogen compounds, polycyclics, and olefins enter into gum-forming reactions. Certain sulfur and nitrogen compounds were found to be most reactive in the gum-forming process. Results of this study indicate that the aromatic constituents of gasolines may contribute substantially and directly to gum formation. A mixture of tetralin and indan-type compounds is reactive, and this mixture appears to be the main aromatic portion entering into gum formation. A secondary but significant amount of naphthalenes also appears to have reacted in the gum-forming process. Analyses of the sulfm compounds in a fresh and an aged fuel indicated that elemental sulfur, hydrogen sulfide, thiols, sulfides, and disulfides were depleted during aging, but thiophenes and residual sulfur compounds were not depleted by fuel deterioration.

Jan 1, 1964

By Edmund Newton

This paper is one of a series comprising the Mineral Industries Survey of the United States, which was begun in 1935. It describes the simple methods of mining and beneficiating manganese ore, in eastern Tennessee and compares these methods with those used elsewhere in the Appalachian region. The report contains general information on manganese to give a clearer understanding of the methods and costs of mining and. beneficiating manganese ore foam a widespread deposit, an in addition presents zone difficulties normally fond in these operations. The marketing, uses, and value of manganese ore are discussed as well as the amount of capital necessary for effective operation on given scale of production, and requirements are outlines for successful operation of manganese, property in this region.

Jan 1, 1941

By Arnis Mangolds

A study of the technical and economic feasibility of using a unique undercut and fill system developed by USBM engineers to mine steeply pitching coal is discussed. A brief resource assessment has indicated that significant reserves exist in most coal regions to justify the development of a safe and economical method of extracting steeply pitching coal. The undercut and fill system combines hydraulic coal cutting with unassisted gravity fluming of the coal slurry. Mining occurs across strike and the extended entry is backfilled to provide uniform strata control and an impermeable water and air barrier. Subsequent mining occurs beneath the previously filled stope such that the roof consists of a solid, controlled back. In addition to providing positive roof and hanging-wall support, the backfill system eliminates subsidence potential, post-mining drainage, and the danger of spontaneous combustion from naturally drafting air currents. A unique through-flowing ventilation system permits passage of return air through the production face via ducting located at the backfill-coal junction. This provides a constant supply of fresh air at the face, eliminates the need for brattice cloths, and provides an additional escapeway for the miners. The preliminary estimates indicate a production rate in the neighborhood of 31.9 tons/ manshift. Capital costs are approximately $3.17/raw ton over a 20-year life with an initial investment per annual ton of $48.29. Total cash-costs for labor, supplies, power and benefits range in the vicinity of $19.59 for a 704,460 ton/year mine, or a total cost of $22.76/ton. These costs do not include federal and local taxes or post-of-money; however, the cost of coal is well within the area selling price.

Jan 1, 1981

By Ernest A. Hersam

"INTRODUCTION.Luring the year 1923 a complete survey was made of the condition of research in the reduction (milling) and concentration of ores and nonmetallic minerals - that is, up to the stage where the ores or concentrates were ready for hydro- or pyro-metallurgical treatment. This survey was carried on under the general supervision of L. A. Lyon, assistant director and chief metallurgist of the Bureau of mines, in cooperation with the milling committee of the American Institute of mining and metallurgical Engineers, the Massachusetts Institute of Technology, and the University of California. Other institutions, including the National Research council, were of assistance and outlined recommendations during the study.While this report is on ""ore dressing"", it contains short sections on coal and clay. The mechanical handling that these minerals undergo is ""dressing"", so that ore dressing is inclusive of metallic and nonmetallic minerals. ""milling"", in its broadest sense, includes with ore dressing certain other metallurgical treatment which the report is not designed to cover.The survey revealed that considerable experimentation is being carried on great cost to the mining industry, but that the research is in general poorly coordinated while companies and firms are adequately taking care of routine tests and the mechanical features involved in their own operations, no comprehensive study of the fundamental principles of ore reduction is now in progress or contemplated. Many industries, besides mining, are interested in the preparation of materials for final treatment, rendering this investigation of value also to them. This class of fundamental research is neglected by industry unless there is some agency which can bring together at least a majority of this interested in a cooperative effort. It can not be effectively handled by a single company; and, moreover, no company can be expected to bear the whole burden of an investigation to benefit the entire industry or a number of industries hence there is need for concerted study within reasonable limits. Furthermore, co-operation eliminates duplication, which is largely a waste of effort and capital."

Mar 1, 1925

By D Nobes, D Trumm, D Manhire, A OÆSullivan, M W. Milke

Coal seam gas (CSG) exploration and development requires the abstraction of significant amounts of water. This is so because gas desorbtion in coal seams takes place only after aquifer pressure has been reduced by prolonged pumping of aquifer water. CSG waters have a specific geochemical signature which is a product of their formation process. These waters have high bicarbonate, high sodium, low calcium, low magnesium, and very low sulphate concentrations. Additionally, chloride concentrations may be high depending on the coal depositional environment. This particular signature is not only useful for exploration purposes, but it also highlights potential environmental issues that can arise as a consequence of CSG water disposal. Since 2002 L&M Coal Seam Gas Ltd and CRL Energy Ltd, have been involved in exploration and development of CSG in New Zealand. Anticipating disposal of CSG waters as a key issue in CSG development, they have been assessing CSG water quality along with exploration work. Coal seam gas water samples from an exploration well in Maramarua closely follow the geochemical signature associated with CSG waters. This has helped to identify CSG potential, while at the same time assessing the chemical characteristics and water generation processes in the aquifer. Neutral pH and high alkalinity suggest that these waters could be easily managed once the sodium and chloride concentrations are reduced to acceptable levels.

Jan 1, 2005

By Roy S. Nutter

The automation of coal mining must occur in a manner that is economically beneficial to the industry and that enhances the safety of the underground and surface workplaces. The work at WVU over the last five years in applying expert systems and neural network controlled autonomous robotics provides the mining industry with tools that when properly applied should clearly lead to mine-wide management and operational improvements and should also contribute to a higher awareness of safety issues by all work crews. This work includes the development of MMSS (Mine Management Support System) and the detailed development of autonomous robotics utilizing neural networks for mining systems.

Jan 1, 1993

By D Craw, D Falconer

The Morning Star mine is one of numerous historically mined gold-quartz lodes in the Preservation Inlet goldfield of southwest Fiordland. The mine is hosted in Early Ordovician metasedimentary rocks that are composed predominantly of psammitic and pelitic lithologies. These rocks still contain their primary depositional features and are characterised by upper greenschist facies metamorphic grade.   The Morning Star mine lies on the eastern limb of a gently south plunging anticline. A bedding parallel fault-vein and an associated network of minor quartz veins hosts mineralisation. The main fault-vein structure strikes N-NW, dips steeply (~60¦) W, has a strike length of ~350 m, a vertical extent of ~170 m and is up to 2 m in thickness. The internal structure of the fault-vein consists of laminated quartz, massive and fibrous quartz, and silicified breccias/cataclasites. Gold is typically coarse grained (~1 - 2 mm) and associated with pyrite, arsenopyrite, chalcopyrite, tetrahedrite and galena. The geometry of the Morning Star mine is similar in form to that of a saddle reef. Gold-quartz lodes in the Preservation Inlet goldfield are comparable in geometry and host rock to their offset equivalents in the Golden Blocks field of NW Nelson.

Jan 1, 2005

By Marie-Eve Boucher

The slag in iron ore pellets allows wet sintering during firing and leads to suitable bounding. The process of slag formation differs whether it takes place in the core or in the external zone of a pellet. In the core, hematite can be reduced to secondary hematite, depending on the temperature and on the carbon monoxide/dioxide partial pressure ratio. The reaction rim that appears at the outskirts of a silica particle in contact with the FeO bearing magnetite is considered as the slag precursor in the core of the pellet. By contrast, in the external zone of a pellet, where O2 is available and CO partial pressure is very low, the first step towards slag formation is flux decomposition, mainly dolomite and limestone. The CaO liberated through this reaction will be available to react with hematite by solid diffusion to form calcioferrites, which play an active part in the slag evolution process. These two slag formation paths will be explained.

May 1, 2007

By G. L. Oldright

The Metallurgical Branch of the Bureau of Mines, as a part of its duties under the Organic Act, works to prevent wastes in industry and to disuse inmate information on the fundamentals of metallurgical processes. During the war, the necessity of large production was increasingly emphasized. Such duties necessitate from time to time the study of units of full commercial size, the operation of which is beyond the facilities of the Branch. When developments have reached this stage; work has been continued in cooperative agreements with operating companies that have been willing to contribute the use of their plants and staffs and to allow the publication of the results obtained for the common good of the industry.

Jan 1, 1948

By Brian H. Greene

The U.S. Army Corps of Engineers, Pittsburgh District, is in the final phases of modernizing the locks and dams on the Monongahela River. One facet of this modernization is to replace the existing Locks 4 located on the Monongahela River approximately 40 river miles upstream from Pittsburgh, Pennsylvania. The existing locks consist of a 56-foot x 720-foot main chamber and a riverside 56-foot x 360-foot auxiliary chamber built in the 1930s. Two new lock chambers, each 84-foot wide by 720-foot long, will be constructed at the location of existing Locks 4 and will be named Charleroi Locks. The purpose of the new project is to facilitate navigation on the lower Monongahela River. Pittsburgh is currently the second largest inland waterway port in the nation, and the new locks will be an important component of the navigational infrastructure. The existing lock walls are of mass concrete construction and are founded on wooden timber piles driven to rock through alluvial sands and gravels. The new lock walls were designed to be supported by approximately 300 reinforced concrete drilled shafts, varying in diameter from 4 to 6 feet. To handle axial loads, the drilled shafts were designed for end bearing, with some skin friction. The tips of the shafts will be founded within competent rock. Since the walls will undergo lateral loads, the drilled shaft design also had to consider this loading condition. Some of the drilled shafts will be installed within coffer boxes, which will incorporate a tremie concrete seal down to top of rock. Complicating the design even further was the fact that some portions of the new walls will be supported by shafts that extend some 9 feet above top of rock through alluvial soils, producing a cantilever effect. The first phase of the new lock construction was initiated in the summer of 2005 with a contract awarded to replace the river wall. This required shutting down the existing riverside lock chamber and diverting all river traffic to the longer existing land chamber. The foundation of the new river wall will use 144 drilled shafts. To optimize the design embedment of drilled shafts in rock for the river wall, an exploratory drilling program was conducted by the contractor in July and August, 2005. Based on the quality of the sedimentary bedrock, the shafts will be embedded in rock from 31 to 38 feet. To achieve adequate end bearing resistance, the tips of each shaft will extend through predominantly claystones into a more competent siltstone unit for a minimum embedment of one foot. The contractor also performed an exploratory pile driving program to aid in the construction of the drilled shaft casing and the cut off walls that will encapsulate the river wall. Two demonstration drilled shafts located outside the footprint of the new river wall, will be constructed prior to the drilling of any production shafts. The demonstration drilled shafts will verify that the contractor is able to meet the design criteria and quality requirements of the contract. The river wall contract is currently underway and is scheduled to be completed in 2010.

Jan 1, 2006

By L. A. Panek, F. D. Wang, M. C. Sun

A method of stability analysis of underground opening by the limiting equilibrium approach is presented. The Coulomb shear failure criterion is employed to find possible fracture surfaces and to evaluate the shear strengths along them. The normal stresses required for the calculation of shear strength, and the shear stresses for the calculation of total shear force along the fracture surfaces, are obtained by the finite-element method of plane-strain analysis. The factor of safety against Coulomb shear failure is calculated as the ratio of the shear strength to the shear force along a fracture surface, similar to the practice in slope stability analysis. Four shapes of openings are analyzed, with three ratios of horizontal lithostatic stress and four coefficients of internal friction. The critical fracture surfaces found by this method of analysis are consistent with the arching of roofs and slabbing of ribs commonly observed.

Jan 1, 1972

By Delon Hampton, Francis K. Mensah-Dwumah

INTRODUCTION The state-of-the art in the area of soil-structure interaction is not sufficiently developed that one can accurately predict the earth load distribution on buried structures. This lack of knowledge has led to the use of high factors of safety in their design and conservative assumptions for the loading diagrams. As a step towards the improvement of the design methodology for buried structures, it is important to be able to apply new analytical techniques and monitor the performance data obtained on full-size buried structures. The performance data should preferably include: 1) the actual stress distribution on the buried structure, 2) stress distribution in the buried structure and surrounding soil media, 3) vertical and lateral movements in the soil adjacent to the buried structure, and 4) deflection of the buried structure. Brown (1967), Davis (1972), Krizek, et al (1974), Krizek and MacQuade (1978), and some few others have compared the performance of full-size culverts with finite element results. Despite all these efforts, a complete understanding of the soil-structure interaction problem has not yet been fully achieved. In an effort to make improvements in the existing design procedures for buried structures constructed in an open-cut, the Department of Environmental Services (DES), Washington, D.C., in conjunction with Howard University, undertook a detailed study of Section II of the Capitol Hill Relief Sewer (CHRS). Inclinometers were used to measure lateral movements of the excavation wall and the total normal interface stresses acting on the concrete sewer were obtained using glotzl total pressure cells. This paper describes the results of the field instrumentation program for acquisition of data, their comparison with the design assumptions, and the results of a series of post-construction finite element analyses of the CHRS project.

Jan 1, 1979

Face and pillar bursts, bumps and bounces are violent failures that occur in underground coal mines in response to a complicated interplay of face and pillar geometry, seam depth, coal properties and interactions between roof, seam and floor strata. Additional complications arise from the presence of gas, mainly methane, and associated pressure and flow that vary with time and are further influenced by the rate of face advance. Where gas plays a dominant role, the term outburst is often used; where stress concentration is primary, terms such as burst, bump or bounce may be used. Hargraves [1983] presents an often cited review of outburst phenomena with emphasis on Australian conditions. Beamish and Crosdale [1998] indicate outbursts have occurred for over a hundred years and present tabulations of outburst numbers in major coal basins throughout the world. A central consideration is the rate of depressurization that occurs in the course of mining. Outbursts are not limited to coal seams and adjacent strata; they also occur in salt and potash mines [Molinda et al 1988]. Stress concentration and gas pressure act in consort, so the combined action is important to understanding the mechanics of bursts or outbursts and to the development of measures to reduce the associated hazard. Gas flow is a complicated phenomenon associated with diffusion of gas from micro-pores and the flow of gas along cleats. Laboratory measurements indicate both are dependent on stress and pressure [Harpalani and Ouyang 1998]. Moreover, gas permeability in porous media is often many times water permeability [Klinkenberg 1941]. However, there is also evidence that gas and water permeability are similar at elevated stress and follow Darcy?s law [Dabous et al]. Gas flow is also of importance to mine ventilation. Price et al [1973] in cooperation with personnel of the former U.S. Bureau of Mines developed finite difference models to estimate methane emission in mines in the Pocahontas No. 3 and Pittsburg seams using a premining pressure of 4.69 MPa (680 psi) and Darcy's law implying laminar flow. However, the effects of stress were not taken into account. In this regard, mine measurements of permeability are usually based on Darcy?s law [Peide 1990]. Early models of gas flow with stress effects were one-dimensional in space and time (x,t) and were developed with the goal of estimating the gas pressure profile from the face into the seam as time passed [Schlanger and Paterson 1987, Zou and Yu, 1999]. A zone of high permeability adjacent to the mining face was recognized in these early models. Later models based on the popular finite element method allowed two spatial dimensions (x,y) and variation in time (t) and for coal seam heterogeneity [Tang et al, 2002]. Zhu et al [2007] formulated a set of nonlinear equations involving a complicated stress-pressure-permeability relationship for application to coal seam gas flow and concluded the coupling was important but failure of coal should also be considered. Wold et al [2008] used a hybrid finite difference - finite element model interleaved with a reservoir simulation model to study outbursts based on extensive mine and laboratory measurements. An important feature of the modeling was the use of spatially variable material properties and Monte Carlo simulation to obtain quantitative estimates of outburst conditions and probabilities. Connell [2009] used the same reservoir simulation model as Wold but used the popular finite difference code FLAC3D for hydro-mechanical responses to evaluate two permeability relationships, one depending on porosity and the other on stress, and concluded that permeability changes with gas production were more complex than computed by either permeability model. Gadde [2010] in an integrated laboratory and field study of coal pillar strength explored the effect of spatial variability and concluded that design based on average strength was acceptable in some circumstances, although gas pressure was not an issue. Liu et al [2011] also applied a FLAC code to analyze two-dimensional gas flow and deformation of layered but homogeneous strata in a Chinese coal mine using a strain-softening model and concluded that damage and gas pressure were important to strata safety relative to the case without gas.

Feb 27, 2013

By Victor F. Paterno

In December of 2007, a joint venture of J.F Shea, Skanska and Schiavone referred to as S-3II, was awarded a $1.15 billion (? 886,880,000) contract to construct the run-ning tunnels and station structures for the Number 7 Line Extension Project in New York City by the Metropolitan Transportation Authority Capital Construction (MTACC) and New York City Transit (NYCT). As part of the contract, a technically complex and critical connection of the new tunnel to the existing, operational 7 Line subway was required. This paper describes the innovative design approach that was conceived to carry out the underpinning and excavation works both within the east-west 7 Line subway tunnel and beneath the north-south, 4 track, 2 level Times Square Station for the A, C & E trains. This underpinning provided temporary support to the subway substructure as the grade of the tunnel invert was lowered up to 10 feet (3 meters) into the rock sub-grade to allow the future subway to pass under the basement of the world?s busiest bus terminal. In addition, the creation of a new portal connecting the existing subway tunnel and the subterranean level of the Port Authority Bus Terminal by means of underpinning structural footings and horizontal spiling/poling of existing ground will be discussed.

By R. Bruce Ferguson

"IntroductionThe Beaverlodge Operation of Eldorado Nuclear Limited is a uranium property located near the north shore of Lake Athabasca about 7 miles from Uranium City, Saskatchewan.The mill is a 2000 tpd carbonate leach operation producing a high grade yellowcake concentrate.The Lamella Thickener was purchased to replace a 55 foot conventional unit which had deteriorated beyond repair. The cost of the unit was considerably lower than for a conventional unit, installation was much simpler and the unit is much more compact allowing it to be installed in the mill basement and be insulated completely.The unit was sized according to the results of tests performed both by the supplier and Beaverlodge engineering staff. The supplier guaranteed overflow clarities of less than 100 ppm suspended solids. The Lamella thickener was developed by A. Johnson Co. of Stockholm, Sweden, is manufactured under license by Parkson Corporation of Ft. Lauderdale, Florida and is distributed in Canada by A. Johnson Co. of Montreal."

Jan 1, 1975