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By Alan A. Campoli

The U.S. Bureau of Mines evaluated two different conventionallongwall gate entry systems in the Southern Appalachian Basin, where mining and geologic conditions are conducive for coal pillar bumps. These gate entry systems were in a coal mine located in the Pocahontas No.3 Coalbed under approximately 2,000 ft of overburden and a massive quartzite sandstone member, and both employed a center abutment pillar flanked by yield pillars. The 8O-ft-square tailgate abutment pillars within the first study area began bumping 500 ft in advance of mining. This was accompanied by face bumps on the tailgate corner of the longwall face. The 120-by 180-ft tailgate abutment pillars within the second study area did not begin to bump until mining was approximately 100 ft past. Coalbed stress change and abutment pillar dilation data demonstrated that a 15-ft-wide perimeter of yielded coal surrounded a highly stressed core in both sizes of abutment pillars. The 62 pet more core area per foot of gate entry in the larger abutment pillars prevented excessive load transfer to the corner of the tailgate panel and eliminated the face bumps experienced with the original gate entry design.

Jan 1, 1990

Pillars in underground limestone mines are required to support the overburden and provide a safe, stable working environment for mining personnel and equipment. Pillar stability was assessed by the National Institute for Occupational Safety and Health in underground limestone mines in the Eastern and Midwestern United States. It was found that current mine layouts have been successful in providing support to the overburden, while a small number of isolated pillar failures were observed. The stable pillar layouts and failed pillars were plotted on a chart that demonstrates the relationship between pillar width-to-height ratio and pillar stress. The results show that pillar failures occurred at the lower range of width-to-height ratios and can occur at relatively low stress levels. Zones indicating the potential risk of pillar instability are shown on the chart, based on hazards associated with the onset of rib spalling, large angular discontinuities and unconfined pillars. The chart can assist limestone mine planners to evaluate potential instability in current or new pillar layouts that are similar to those included on the stability chart.

Jan 1, 2008

By L. L. Oden

The Bureau of Mines analyzed vapor-aluminum-diffused stainless steels prior to conducting corrosion measurements in severely sulfidizing environments. These studies were conducted in response to the national need to better control the environmental and economic costs of corrosion. Specimens of a ferritic alloy and austenitic alloys types 316 and 310 were treated by a proprietary vapor-aluminum diffusion process and a dual vapor-aluminum diffusion process. Weights and dimensions were recorded before and after treatment, and comprehensive metallographic, X-ray, and electron micro-probe analyses were conducted for both single-and dual-vapor-aluminum-diffused specimens. Surface aluminum concentrations ranged from 26 to 34 wt-pct for all alloys. Dual treatment increased the quantity of aluminum per unit area by a factor of 2 and increased the depth of penetration by a factor of 3. The aluminum-diffused region of the ferritic alloy was single-phase aFe solid solution that was ordered on the surface, whereas two distinct diffusion layers were ob-served on the austenitic alloys. The outer layer comprised aFe solid solution that was ordered on the surface, and the inner layer was a two-phase mixture of aFe solid solution and the interrnetallic compound AINi.

Jan 1, 1982

By R. W. Bartlett

Hydrometallurgical processes for copper flotation concentrates avoid the SO2 emission problems associated with smelting, but they require oxidation of copper to obtain solubilization during leaching. The resulting copper ions must be reduced to metal, which normally requires large amounts of energy or energy equivalent reagents such as hydrogen or iron. The conventional reduction method is electrowinning which is labor intensive and involves 35 to 40% of both the entire operating and capital cost of the better hydrometallurgical processes. There is considerable incentive to reduce the cost and energy required for producing copper ions from solution. Furthermore, it is thermodynamically possible to do this using sulfur contained in the concentrate itself. In fact, the thermodynamic potential of oxidizing sulfur from S+4 to the S+6 state is greater than required to reduce cupric ion to metallic copper.1 Realization of this potential in a practical process would significantly reduce the cost of hydrometallurgical processing of sulfide concentrates because sulfur would become a cost-free reducing agent. An apparently practical method is to use SO: and HSO3 ions, obtained by ammonia scrubbing of SO2 from a concentrate roaster gas, to reduce leached copper to metal in an autoclave. This reduction step is the Jumau 1907 A series of steps, including sulfite reduction, constituting a crude flowsheet was suggested by Parker in 1976.3 The present paper describes a detailed complete. sulfite process flowsheet based on materials and energy balances, discusses the chemistry, and discusses further work that is required to develop this process. Process materials, utilities costs, and energy requirements are projected to be unusually low because of the utilization of sulfur from the concentrate. Process Description The sulfite reduction process requires a copper concentrate containing approximately as much sulfur on a weight basis as copper. For those few concentrates that are deficient in sulfur, additional sulfur can generally be obtained by allowing more pyrite to float in the concentration process. The process steps, shown in [Fig. 1], are (1) a fluid bed roast of the concentrate, (2) a leach of the roaster calcine using recycled barren liquor from the reduction autoclave, (3) scrubbing of the SO2 gas with ammonium sulfite liquor to produce ammonium bisulfite, (4) reduction and precipitation of copper in an autoclave by combining part of the SO2 scrubber effluent (SO5

Jan 1, 1980

By G. R. Mansfield, W. B. Lang

THE third year of Government exploration for potash by the U. S. Geological Survey and the U. S. Bureau of Mines under the authorization of the act approved June 25, 1926 (Public 424-69th Cong.) is drawing to a close. The results thus far obtained are thought to be sufficiently interesting and satisfactory to justify the presentation of a brief preliminary account of the work. Review of Earlier Exploration The general geologic similarity of the Permian basin of Texas and New Mexico to the potash field of Germany had long ago suggested the possibility of the presence of potash salts somewhere in that basin, but petroleum exploration was then making little headway there and the necessary data could not be obtained. The discovery by Udden in 1912 of potash in brine from the Spur well in Dickens County, Texas, was the first direct evidence of its presence. In 1915-17 the U. S. Geological Survey drilled a test well with standard tools at Cliffside near Amarillo, but failed to find potash. Nevertheless, an observer has been kept in the field, at first in cooperation with the Texas Bureau of Economic Geology but since 1922 by the Survey alone, to secure data and samples from oil and gas tests. Since the discovery of oil in Reagan County in 1923 exploration by drilling has been rapidly extended to many parts of western Texas and southeastern New Mexico. The identification of the potash-bearing mineral polyhalite was first announced by the Geological Survey in June, 1921, the material having come from cuttings from the Bryant well in Midland County and the Burns well in Dawson County, Texas. Successive discoveries and announcements of similar purport aroused widespread interest in potash exploration, which culminated in the action by Congress already noted. Permian Salt Basin The Permian salt basin extends generally northeastward from western Texas and southeastern New Mexico through the Panhandle sections of

Jan 1, 1929

By Isaac A. Zlochower, Michael J. Sapko, Eric S. Weiss, Kenneth L. Cashdollar

This paper describes dust explosion research conducted in an experimental mine and in a 20-L laboratory chamber at the Pittsburgh Research Laboratory (PRL) of the National Institute for Occupational Safety and Health (NIOSH). The primary purpose of this research is to improve safety in mining, but the data are also useful to other industries that manufacture, process, or use combustible dusts. Explosion characteristics such as the minimum explosible concentration and the rock dust inerting requirements were measured for various combustible dusts from the mining industries. These dusts included bituminous coats, gilsonite, oil shales, and sulfide ores. The full-scale tests were conducted in the Lake Lynn experimental mine of NIOSH. The mine tests were initiated by a methane-air explosion at the face (closed end) that both entrained and ignited the dust. The laboratory-scale tests were conducted in the 20-L chamber using ignitors of various energies. One purpose of the laboratory and mine comparison is to determine the conditions under which the laboratory tests best simulate the full-scale tests. The results of this research showed relatively good agreement between the laboratory and the large-scale tests in determining explosion limits. Full-scale experiments in the experimental mine were also conducted to evaluate the explosion resistance characteristics of seals that are used to separate non-ventilated, inactive workings from active workings of a mine. Results of these explosion tests show significant increases in explosion overpressure due to added coal dust and indications of pressure piling.

By Piotr M. Kapias

Keywords: extraction, lead refining, Pb-Tl solutions, PbCl2-TlCl solutions. The theoretical part of the paper presents an analysis of the conditions of Tl extraction from Pb using different extractants and determines the activity coefficients in liquid solutions of Pb-Tl and PbCl2-TlCl. The dependence of the solution activity coefficients upon the composition and temperature is presented in the form of Krupkowski and Krupkowski-Fitzner equations. An analysis and evaluation were made of the influence of the methods of PbCl2 introduction into a Pb-Tl solution upon the efficiency of the Tl extraction process under thermodynamic equilibrium conditions. The analysis made it possible to formulate relevant equations describing the dynamics of the process of Tl extraction from Pb under stirring conditions as well as under the extractant injection into the liquid Pb-Tl solution conditions. In the part concerning experimental work, results are given for tests on the Tl extraction process from Pb-Tl liquid solutions of initial Tl content of approximately 0.05% mass using PbCl2 or, alternatively, Cl2(g) conducted in laboratorial scale. The tests made it possible to establish the influence of time, temperature, mass and the method of introducing the extractants into the liquid upon the level of Tl extraction from a Pb-Tl solution. Labels used in the text. WT1m - TI content in metallic phase, % mass WT1s - TI content in chloride phase, % mass wpTlm - TI content in metallic phase in initial state, % mass W°Tlm - TI content in metallic phase in equilibrium conditions, % mass m°m - initial mass in metallic phase mm - mass in metallic phase mPbC12 - lead chloride mass fed to the system MPbC12-TIC1 - mass in chloride phase MPb - lead atomic mass, g/mole MT1 - thallium atomic mass, g/mole K - Nernst distribution coefficient Xi - mole friction of solution component X°PbC12 - mole friction of PbC12 in saturated solution X°T1C1 - mole friction of T1C1 in saturated solution T - temperature, K R - gas constant, J/(mole K) k - kinetic equation constant, 1/(%-min.) ?i - solution component activity coefficient e - extraction rate, %

Jan 1, 2008

Some coal was mined in Wyoming for use at the forts as early as 1859, but the commercial development began in 1867, and mining increased with the completion of the Union Pacific Railroad. Tonnages produced are shown in Table 78. [ ]

Jan 1, 1942

By Andrew Bamber

Between the early 20th Century and the present, the production of mined copper ore has increased from tens of millions of tonnes per annum to several billion tonnes per annum, an unprecedented increase. Contemporaneously, the average grade of mined copper ore has declined from over 2% to approximately 0.5%. Again, in parallel, the price of copper has increased dramatically to reflect the increasing cost of extracting and beneficiating such large amounts of low-grade ore. Within the period the copper price however exhibited several periods of intense volatility, driven by interrelationships between demand and supply, themselves strongly driven by periodic industry developments in the fields of exploration, ore extraction, mineral processing and extractive metallurgy. It is notable that these periods of intense innovation exhibit a strong correlation to preceding periods of price, demand or production volatility. This paper examines key trends in each theme between 1900 and 2018 and draws strong conclusions as to the interrelationship between them. INTRODUCTION The production of copper metal has increased from around half a million tonnes per annum to over fifteen million tonnes per annum over the course of the 20th Century. Contemporaneously, the average grade of copper ore mined has declined from over 2% to approximately 0.5% copper by weight. As a result, the volume of Run-of-Mine (ROM) copper ore required has increased from around ten million tonnes per annum to over a five billion tonnes per annum over the period, an unprecedented increase. The adage is that demand drives supply, that prices rise in times of high demand and/or short supply, and that prices fall when supply catches up with demand. The market sellers in biblical times knew this well; mediaeval market sellers equally. Dutch tulip sellers in the 1600’s experienced both in extremis — much to their dismay — and the great industrialists and early capitalists of the industrial revolution knew this also. Mining recently went through a cycle where investors and producers both believed that the “old” rules did not apply any more and that the industry was in a “supercycle” where prices would continue to rise on the back of increasing demand. However, as the GFC of 2008–2011 showed, this was not to be. Another adage says that we constantly innovate to lower costs of production. Is this true? Do we only innovate to lower cost? Has the true cost of producing copper actually gone down over the century? Or are there other balancing forces at play? Herb Kellogg, in his Centenary address to the American Institute of Mining Engineers (now SME), categorized two types of innovation: ‘Bigger and Better’, and ‘New Process’ (Kellogg, 1970). ‘Bigger and Better’ allows increased production at presumably lower cost; ‘New Process’ supports the same result at presumably the same scale, but at significantly enhanced efficiencies and (therefore) lower cost. It is suggested that both types of innovation are constantly and interchangeably required to achieve the kind of scale-up, at achievable operating and capital cost that the industry has achieved over the period. What can be shown, over this last Century and a bit, is that supply and demand are indeed inextricably linked, price and demand are linked, but so are price and cost of production. And the innovation required to sustain that weaves in between them all, cost-effectively increasing production capacity or lowering the cost of production at the same throughput, and sometimes both.

Jan 1, 2019

The Goonumbla district of central-west New South Wales, Australia, hosts four significant porphyry Cu-Au deposits that contain a combined probable ore reserve of 71.7 million tonnes at 1.2 per cent Cu and 0.5 g/t Au. Mineralisation in the Goonumbla deposits is associated with narrow, pipe-like intrusions of quartz monzonite porphyry, in which several intrusive phases have been recognised. Five textural characteristics have been used to distinguish the various intrusive phases recognised in the Goonumbla porphyry deposits: (1) phenocryst abundance; (2) nature of the groundmass; (3 and 4) proportion and mineralogy of the mafic and non-mafic phenocrysts; and (5) the presence or absence of anhedral quartz grains in the groundmass. In addition, each intrusive phase is classified according to its estimated original quartz, alkali feldspar and plagioclase contents. Features common to all of the intrusions are that they contain two types of plagioclase and two or three types of K-feldspar phenocrysts in variable proportions. Most of the intrusions also contain apatite micro-phenocrysts. The groundmass textures range from aphanitic through granular to euhedral K-feldspar in a quartz matrix. With only one exception, all of the intrusive phases recognised are either quartz monzonite porphyry or mafic (>5 per cent mafic minerals) quartz monzonite porphyry. The exception is a biotite - phyric monzodiorite, which occurs in the basal portions of one of the deposits. Textural variants of intrusive phases at Goonumbla have been investigated to establish whether specific combinations of characteristics indicate mineralised intrusions, while others suggest barren ones. Some correlations between intrusion types and Cu and Au grade have been recognised. The most important textural characteristic that relates to high grade mineralisation appears to be the low proportion of mafic minerals, because the quartz monzonite porphyry intrusions hosting most of the mineralisation in the two deposits studied contain

Jan 1, 1999

By A. Satter

Studies of wellbore heat transtnission during the injection of a hot fluid, as either gas or liquid, have appeared in he literature. The present investigation takes into account the effect of condensation, which is of practical sig-nificance when considering steam injection operations. Calculation procedures are given for superheated and for saturated and undersaturated steam. Effects of injection rate, time, pressure, temperature and depth of a well on heat losses have been analyzed. The benefits of using a packer are discussed. Also presented are correlations for estimating heat losses involved during injection of saturated steam. INTRODUCTION Studies of wellbore heat losses during the injection of a hot fluid have appeared in the literature. Ramey' has presented a useful solution for calculating temperature of a heated gas as a function of depth and injection time. Ramey and others have also presented solutions for the hot liquid case.'. ' No method is available to take into account the effect of condensation, which is of practical significance when considering steam injection operations. This paper presents a method of estimating the quality of a condensing fluid as a function of depth and time. The approach is basically the same as that used by Ramey. The following injected fluids and conditions are considered in this study: 1. Injection of superheated steam A. Decrease in temperature until cooled to the saturated state B. Condensation until condensed completely into hot (saturated) water 2. Injection of saturated or undersaturated steam; condensation until condensed completely into hot (saturated) water. Although this investigation does not include cooling of water after steam is completely condensed, Ramey's solution for the hot liquid case can be used for this purpose. The effects of various process variables on heat losses During the injection of steam into a well are investigated. THEORY QUALITY OF CONDENSING STEAM Predicting the behavior of condensing steam flowing down the injection wellbore requires an estimate of the quality of steam, i.e., the mass fraction of vapor in the mixture. Let us divide the total depth of the injection well into several segments which might be, but do not have to be, of equal magnitude. Consider condensation within a given depth interval AZ, the bottom of which is located at a depth Z from the surface (Fig 1). The following equation, derived in the Appendix- relates the quality of steam at the bottom of an interval to that at the top of the interval:

Jan 1, 1966

By K. A. Natarajan, M. K. Yelloji Rao

The electrochemistry of sphalerite and chalcopyrite was investigated under both open-circuit and applied potential conditions. Emphasis was placed on the selective dissolution of sphalerite from a mixture of sphalerite-chalcopyrite flotation concentrates as a function of applied potential, acid concentration, pulp density, and time. Morphological studies using a scanning electron microscope (SEM) were also carried out with chemically leached, galvanically leached, and unleached surfaces of sphalerite and chalcopyrite. Selective dissoulution of sphalerite could be achieved if the mixed chalcopyrite-sphalerite concentrate slurry was maintained at a negative applied potential of -0.4V (SCE). Probable mechanisms involved in the selective dissolution of zinc are disscussed.

Jan 1, 1990

By C. F. Austin, J. K. Pringle

In view of the many large differences in structure and physical properties between metals, plastics, and most rock materials, a continuing experimental petrodynamics program has been established at the U. S. Naval Ordnance Test Station. Part of this program has been to test the concepts of scabbing and corner fracturing in a typical coarse-grained rock. The results of this experimental program have shown very interesting differences between the formation of fractures in rock and the formation of fractures in other materials such as metals and plastics which are often used as models for rock behavior studies. The gross behavior of coarse-grained brittle solids subjected to impulsive loading has undergone abundant empirical investigation by mining organizations and by mineral research agencies. However, the fundamental response of coarse-grained brittle solids to impulsive loading has received little detailed study in recent years. The various organizations and agencies interested in rock-like brittle materials have broadly applied the stress wave interaction concepts of the impulsive failure of metallic samples to the behavior of granular brittle materials 1,2,3,4,5,6 Using the metallic sample approach for stress wave induced fracturing, an understanding of the fracture behavior of brittle materials which perhaps is not always appropriate has been established in the open literature. An excellent summary of the metallic sample approach is provided by ref. 7 which summarizes the general beliefs prevalent at present regarding the impulsive behavior of brittle solids: "In the blasting process, rock breakage occurs primarily from tensile slabbing, beginning at a free face and proceeding back toward the explosive charge. The tensile waves formed by the reflection of incident compressive waves from two or more free faces will collide in the rock, resulting in large localized tensile stresses. Fractures formed by the interaction of these tensile stresses begin at the apex of the free face planes and progress back into the rock as shown...."

Jan 1, 1964

By Maurice Deul

A special short-term research project was initiated to develop plant design parameters for treating mine waters to yield an effluent acceptable for discharge into streams and a sludge amenable to effective and economic disposal. Neutralization was investigated to provide simple and direct treatment of ferruginous acid waters using lime; coarse limestone, and limestone followed by lime. Limestone treatment was accomplished using a small cement mixer as a simple reactor to provide abrasive agitation of limestone and mine water. Results from tests with nine mine waters encompassing a wide range of iron and acid concentrations showed that coarse limestone, one of the cheapest neutralizing agents known, is potentially useful for treating mine water discharges. However, process variables must be defined more completely before the practical applications and limitations of the process can be fully established. In all the tests a pH of >7.2 was obtained ultimately with limestone. Reaction time to reduce the iron content below 7 ppm in the water was dependent on the original ferrous iron concentration. For waters with high ferrous iron content, treatment with limestone followed by lime rapidly produced an acceptable treated water, Sludges with good settling and handling characteristics resulted when limestone was used for neutralization either alone or with subsequent treatment. Reaction rates are expressed graphically as pH change and iron depletion curves; characteristic solids settling curves are shown. The temperature dependency of ferrous iron oxidation was verified.

Jan 1, 1967

By W. Stumpf

Mill log analyses and complementary hot deformation studies were carried out on the as-cast structure and on a hot rolled structure of the same low carbon steel (SAE 1006) that had been processed according to the direct or hot charge route in a compact strip plant. The main softening mechanisms during hot rolling were found to be complete dynamic recrystallization during each pass above about 1000°C for the as-cast structure with an austenite grain size of 277µm and, secondly, incomplete static recrystallization after each pass with a finer austenite grain size of about 26 µm and tested below 1000°C. A two-stage hot working mechanism has also been found by others on a similar steel also processed by the hot charge route and this may appear to be a characteristic of this type of material. Constitutive constants determined from hot working studies were combined with other constants from the literature and incorporated into a grain size development model that correctly predicts the final ferrite grain size in the hot rolled strip steel. The predicted incomplete static recrystallization during the last three finishing passes in the plant appears to result in some retained strain in the hot rolled austenite as it transforms to ferrite on the run-out table. Analyses with the model predict relatively little sensitivity of the final ferrite grain size to process and product variables such as temperature and strain per pass for the last three passes and also the starting austenite grain size in the slab. Keywords: Low carbon steel, hot rolling of strip steel, rolling mill log analyses, austenite grain size, ferrite grain size, grain size modelling.

Jan 1, 2003

On 26 August 1971, the Bureau of Mines, through the Spokane Mining Research Center at Spokane, Washington, approved initiation of the evaluation of oscillatory energy toward enhancing flow of polymer into shale. The effort was an outgrowth of work performed under Bureau of Mines Contract H0210055, "Feasibility Study of Surface Impregnation equipment for the Chemical Stabilization of Coal Mine Structures," wherein feasibility of the impregnation concept was expressed and the detailed design of a machine for impregnation of coal mine roofs with polymer (under static pressure head) accomplished. Under these efforts, the potential enhance¬ment of polymer uptake into the shale thru superposition of oscillatory energy on the static impregnation pressure was suggested. A proposal for investigation of this phenomena entitled "Feasibility Study of the Application of Oscillatory Energy as an Aid to Afford Chemical Stabilization" was presented to the Bureau whereupon a three month program for initiation of investigation was approved.

Jan 1, 1972

By I. Sedat Buyuksagis, Jamal Rostami, Ali Naeimipour

"Improving the safety of underground openings and optimizing ground support systems requires reliable ground characterization. This includes understanding the joints, discontinuities, and rock strength. Joint information can be obtained from borehole cameras, but measuring in-situ rock strength is very difficult. Use of borehole probes can be an ideal and efficient way to meet this requirement. This paper discusses the development of the Rock Strength Borehole Probe (RSBP) and presents the preliminary laboratory/field test. This device can be an accurate, flexible, quick, non-disruptive, and cost-effective alternative to estimate the rock strength inside boreholes in underground mines and tunnels. In addition, the scratch testing procedure in the laboratory by means of a newly developed miniature linear cutting machine is explained. Based on the results of these tests, two equations are developed to estimate the uniaxial compressive strength (UCS) and Brazilian tensile strength (BTS) of the sedimentary/metamorphic rocks. These equations, are based on testing 27 different sedimentary, igneous, and metamorphic rocks by full-scale scratch tests, including the cutting tests by a miniature disc. The results show a good correlation between the normal force and the compressive strength of sedimentary rock if the depth of the scratch is known. No significant correlation was observed for igneous rocks, due to the impacts of grain size.INTRODUCTIONEstimated rock mass properties are essential input for geomechanical studies and stability analysis for any structure constructed in or with rock. The availability of geological information is the crucial component of rock mass classification. One of the important parameters in evaluating rock mass properties is the intact rock strength. This parameter is usually measured by testing the core samples, obtained from exploration boreholes. These samples are subsequently tested in geomechanics laboratories. The test results offer limited information about the rock at the few boring locations despite all the efforts and costs. In addition, these results may not necessarily be representative of the behavior of the rock in the field as the test cannot provide the insitu condition of the ground."

Jan 1, 2016

By Kenneth E. Hjelmstad

The U.S. Bureau of Mines (USBM) has developed an electromagnetic (EM) fire warning alarm sys-tem for underground mines. The system generates a magnetic field for through-the-earth signal trans-mission to microreceivers carried by individual miners. EM fields can induce electric currents in metallic conductors; if an EM transmitting antenna and an electric blasting cap array are too close, the field could induce a current in the blasting array and cause an unintentional initiation of the electric detonators. The USBM conducted tests that define the safe and unsafe regions for using electric detonators near the transmitting antenna of the warning alarm system. The minimum safe distance between a transmitting antenna and a blasting array is the distance where the induced electrical current in the blasting array is 50 mA, which is the safe current level specified in Federal mine safety regulations. Tests indicate that at transmitting power levels of 100-and 1,000-W, separation distances of 9 and 21 m, respectively, were required. These distances are small compared with the dimensions of a mine. Thus, the tests indicate that, with proper placement of the transmitting antenna, the warning system can be used safely in the proximity of blasting arrays.

Jan 1, 2010

By Manuel Gomez

The heat contents of thermally dried lignite, subbituminous B coal, and their corresponding low-temperature chars were determined at various temperatures below the highest temperature to which these materials had been exposed before testing. Similar studies were made on high-temperature char and coke, spent shale, and ash prepared from both coal and spent shale. Heat content of the carbonaceous materials tested correlated well with temperature. Heat con-tent results for spent shale compared well with the published results. Statistical analysis of experimental data together with published information showed moisture, volatile matter, carbon content, and temperature to be the principal variables influencing the specific heat of coal. The specific heat of devolatilized carbonaceous substances, char and coke, approached that of graphite as the carbonization temperature was increased. The elemental specific heats of coal constituents, estimated from the ultimate analysis and the specific heat of coal, compared favorably with published elemental specific heats.

Jan 1, 1965

By Vivian M. Troughton, Clifford J. Wren, Tony P. Suckling

"Secant pile walls are used for the construction of deep basement walls and earth retaining structures. Risk associated with secant pile walls is commonly perceived as being the safety of site personnel and the public during piling operations, or safety used in design. It is recognized that both of these are of paramount importance, but additional risks associated with the construction methods must also be considered to complete the risk management profile for any project. For the particular case of secant pile walls, if Clients take a ""back seat"" they may be unaware of the total risk package they are procuring. The authors argue that the piling contractor should be involved at the earliest stage giving input into the overall risk management process. Such early involvement within the professional team will maximise cost benefits to the Client and minimise the delays and cost increases sometimes associated with these works. A consistent methodology for identifying and then managing all the risks associated with secant pile walls is proposed. An example is used to demonstrate how this methodology should be employed. INTRODUCTIONThis paper discusses the value of a complete and integrated risk management process for structures with basements formed using a secant pile wall.The risk associated with any secant pile wall can be consicfered as comprising four components;1. Safety of people on site2. Safety of design or design methodology3. Construction risks4. Commercial risksAny risk management system needs to fully consider and integrate each of these four components to maximise the likelihood of a successful outcome for the basement.This paper concentrates on the construction risks for secant pile walls, as these are not often understood by the Client (in this paper the term ""Client"" also includes the main contractor, principal contractor, Engineer, Planning Supervisor, quantity surveyor and cost consultant)."

Jan 1, 2005