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By Wallace Thorneycroft

This paper by Mr. Thorneycroft, Past President of the Institution of Mining Engineers (Great Britain), and chairman of its Subsidence Committee, is a valuable contribution to the assemblage of data on ground movement and subsidence from coal mining. It deals with the effects of advancing longwall, a method almost universal in Europe but as yet little used in the United States outside of Northern Illinois and in a few scattered places elsewhere in thin coal beds, as in the Lexington field, Missouri. Canon City field, Colorado, and the Birmingham field, Alabama. Although by advancing lougwall, a maximum recovery of coal is obtained and it has to be employed where there are weak shale roofs, especially in deep mines, it is a more expensive method of operation unless the mining conditions are suitable, which is unusual in the United States. On the other hand, there is a growing use in this country of what is essentially retreating longwall, with an approximately straight break-line as described by Rayburn.1 With this method, in which the pillars are substantial, it was shown that the ground movement effect is practically identical with that of advancing longwall, hence the value to mining engineers of the precise data given by Mr. Thorneycroft. His study is unusual in several respects; it deals with conditions connected with the mining of a coal bed 20 in. thick, about 490 ft. below the surface, over a mined-out area in a 5-ft. bed about 90 ft. below, which had been mined about 12 years earlier. Both mines passed under Plean House, a building of massive cut stone construction built over a century ago, in which Mr. Thorneycroft resided throughout the under-mining period. Slight cracks opened in the walls as the longwall face approached the house in mining the upper seam, and closed after the face passed and subsidence took place. In the northwest wall, the last wall the face passed under, some stone blocks were broken, which was attributed to a strike period when the advance of face came nearly to a standstill. Other than this there was no perceptible permanent damage, as indicated by the photographs taken by the undersigned in 1923 (Figs. 1 and 2). Mr. Thorneycroft used a novel means of studying the successive positions of the corners of the building: he hung plumb bobs from the eaves and traced the paths of the respective plumb bob points on level surfaces. The angles made by the house walls with the plumb bob lines gave measures of the slope of the subsidence wave at the respective points. A slight upward wave was indicated by the surveys ahead of the subsidence, which would appear to be due to a tilting over of the massive sandstone stratum in the roof measures near the surface, the edge of the coal face acting as the fulcrum.

Jan 1, 1931

By N. I. Fedunets

The development of variousworks on automatising of operational processes at coal¬extractive and coal-mining enterprises introduces an ever growing necessity to carry out investigations in the wide range of problems concerning the possibitity of joint functioning of different levels, which seems impossible without the use of integrated systems of passing on and prossessing the information on the basis of modern calculating machines and other means of automation. This view on the problem of automation of the enterprise operation presupposes that the structure of automated operational dispatcher control (AODC) should be very closely connected with that of integrated data processing system! In this case the AODC structure from the point of view of automation of its functions (or operations), i.e. from the point of view of informational techniques coinsides in general with the structure of dispersed information and calculating network (ICN), realizing the passing on and processing of the information as well as solving the functional tasks on the basis of local calculating networks of electrical calculating machines with correspond to different levels of AODC and located with regard to productional hierarchy of operational dispatcher control (ODC). In this approach however the use of dispersed, ICN gives opportunity to. realize the possibility of action only with realisation on the CM of algorithmic' task of ODC, when a dispatcher 'is the lastto make a final decision.. At the mining enterprises it's rather difficult for a dispatcher to make a final decision because of the great variaty of productional situations, the presence of contradictions and a number of other subjective factors which often lead to the dispatcher's mistaks causing great material losses of the enterprise. Hence the necessity arises to automize and formalize making a decision and including the expert systems (ES) into the ontour of AODC, which enables the dispatcher 2to get a neeessary advice in different situations, especially in case ofemergency on the basis of establishing the mechanism of logical layout of non-formalized data base (DB) procedures relying upon the background knowledge and experience of experts. The presently presented in this paper theory of establishing the. ISAODC on the basis of dispersed ICN and expert systems is aimed at trasformation ofoperational' tasks of the higher level into the technological ones, at solving the tasks of diagnostics and forseeing of the process of production, with a dispatcher getting an opportunity of making qualitative and effective decision concerrning the choice of operational impacts in actual time on the basis of generalization the realisation of presented conception must, as has been already mentioned, be accomplished only on the basis of the foundatoin of the united dispersed informational - calculating network of electrical calculating machines.. This includes ICN which would be operating in . different operational contours. . From what has just been said it followsthat the ISAODC Structure will conside with that of DICN (in our case) and hence all the problems of ISAODC structure would be regarded as the questions of DICN.

Jan 1, 1992

By W. Jacobi, F. Steinhausler, A. C. James

INTRODUCTION In 1979 the Committee on Radiological Protection and Public Health of the Nuclear Energy Agency (OECD) defined a programme of work needed to improve radon dosimetry and monitoring. A Group of Experts in this field was established by the NEA and at the first meeting of the Group the authors were designated an initial task with the following terms of reference:- To review dosimetry of radon, thoron and their daughters, and in particular:- - to examine analytical models for the assessment of dose to individuals who are exposed to aerosol mixtures of these nuclides; - to examine the state of knowledge of the characteristics of aerosol mixtures of these nuclides and other factors that influence the dose to the lung, and identify those parameters which should be measured for dosimetric purposes in radiological surveillance programmes; - to identify areas where further study or research are needed. In this presentation we attempt to examine, with these objectives, the relationship between exposure to radon and thoron daughter aerosols expressed in terms of the working level month l (WLM), or its SI equivalent, and dose accumulated by sensitive tissues of the lower respiratory tract. Over the past two decades many attempts have been made to calculate the lung dose equivalent to l WLM exposure in mine atmospheres, usually in terms of the maximum dose to basal cells in a small part of the bronchial tree. The calculations have employed variously detailed models of airway anatomy, aerosol deposition, ciliary clearance and dosimetry of tissue at the bronchial surface, and are well reviewed elsewhere (Walsh, 1970; UNSCEAR, 1977; Jacobi, 1977; Fry, 1977). They agree in the conclusion that basal cells in segmental bronchi (between about the third and fifth bronchial generations) receive the highest dose, although estimates of this dose differ by more than an order of magnitude, ranging from about 0.2 to 12 rads per WLM (2 mGy to 120 mGy). The most critical parameters of these models are the assumed depths of basal cells below the bronchial surface and the fraction of the WL carried by free ions. When more realistic values of these parameters are substituted (Jacobi, 1977), the dispersion in calculated maximum doses to basal cells is reduced to give a much smaller range from about 0.3 to 1 rad/WLM (3 to 10 mGy/WLM). The parameters and methods employed in these various models, however, remain too diverse to infer the sensitivity of the dose conversion to practical variations in aerosol characteristics and the additional factors determining lung deposition and clearance. We have therefore taken the approach of developing newer models (Jacobi and Eisfeld, 1980; James et al, 1980) that can encompass the range of available experimental data on the physical and biological behaviour of radon and thoron daughters and also accommodate the different interpretations that can be put upon them. We have paid particular attention to the averaging of dose over tissues of equal sensitivity in the lung

Jan 1, 1981

By R. W. Gurry

In the course of a series of measurements of the rate of diffusion of carbon in austenite at about 960°C. (1760°F.) and III0°C. (2030°F.), it became necessary to determine carbon concentration when austenite is saturated with graphite, because of indications that the hitherto accepted value of the solubility of graphite in this temperature range is too low. Specimens of carbonyl iron were saturated at temperature in a hydrogen-toluene atmosphere which was in equilibrium with graphite. The carbon conteDt of the steel was measured either by combustion or by the loss of weight after substantially complete decarburization in an atmosphere of hydrogen saturated at room temperature with water vapor. The absence of carbide or graphite as a phase at temperature was demonstrated by microscopic examination of saturated specimens that had been drastically quenched, The mean result of concordant ments-—taeats-is-^that gamma iron saturated with respect to graphite contains 1.39 per cent carbon at 957°C .(i755°F.) and 1.89 per cent at III0°C. (2030°F.). These two values, plotted along with what appears from the literature to be the best value (0.69 per cent) at the iron-graphite eutec-toid temperature (738°C. or I360°F.), yield a curve that is almost linear, and which, continued upward through a short interval, leads to a limiting solubility of 1.98 per cent at the iron-graphite eutectic temperature II35°C. (2075°F.) instead of the commonly accepted value of about 1.7 per cent. Experimental Procedure The samples of carbonyl iron, with the exception of those for microscopic examination, were cylinders 3.8 cm. long, 0.475 cm. in diameter. They were placed in a graphite holder and suspended in a vertical furnace, the temperature of which was controlled precisely by means of the resistance of the platinum winding.' In the empty furnace the range of temperature over a distance of 4 cm. at the heat center was 9°C. (I6°F.) but at a given location the temperature remained within +o.5°C. (o.9°F.); it was measured by inserting a platinum-rhodium couple calibrated at the melting point of palladium, 1555°C. Since the presence of the sample probably improved the uniformity of temperature, it was decided to consider the average temperature of the sample to be 3°C. lower than the maximum measured in the empty furnace with the same setting of the controller. Data were obtained in the vicinity of two temperatures, 960°C. (I760°F.) and III0°C. (2030°F.). The samples for microscopic examination, 2.2 cm. long, 0.475 cm. in diameter, were suspended individually on pure iron wire; for these no temperature correction was applied. The carburizing gas passed into the furnace was dry, oxygen-free hydrogen saturated at room temperature with toluene; this quickly produced on the

Jan 1, 1942

By C. D. Rugen

Ground raw material as fed to the cement kiln generally is a mixture of two to four components, each of which may have widely varying physical and grinda-bility characteristics. Chemically similar materials found in the same or different deposits may also differ appreciably in physical properties, depending on their crystalline structure. The calcareous or principal cement raw material may vary from a soft chalky limestone or friable cement rock at one plant to a hard limestone or oyster shell at another plant. Argillaceous materials include clays, shales and granulated slag. Siliceous and/or iron-bearing material, including sand, iron ore and cinders, may also be required in addition for certain types of mixes. Frequently differences in grindability characteristics are not given the attention they deserve in comparing grinding results at one plant with those at another or in selecting suitable circuits or equipment. Many types of crushing and grinding units are used, with numerous open and closed circuit arrangements. Both wet and dry processes are used successfully. When rebuilding or installing a new grinding department, the choice of process or circuit in some cases may be dictated by the nature of the materials being ground but in other cases it may be the result of good salesmanship or a compromise. An over-all evaluation of wet process vs. dry process grinding should not be made without considering the effects on all other producing departments in the plant. In general, the more recent grinding installations, both wet and dry, have employed closed circuits, but many older open circuits are still in successful operation. In our experience, closed-circuit raw grinding pays, because of lower power consumption and less operating labor per ton of product. However, some dry open-circuit mills with improved equipment and operating methods compare favorably in performance with dry closed-circuit units. Character of Ground Raw Mix (Kiln Feed) Grinding in the cement plant in general requires a considerably her product than in most ore-grinding operations. A distinction between fineness of kiln feed and of cement may be made in that kiln feed preferably should be ground to a limiting size,whereas finished cement should include a range of size fractions including sufficient fine flours to give a product with satisfactory physical qualities. Grinding raw mix for kiln feed therefore is more comparable to ore grinding, although the limiting size is smaller. Most kiln feeds are ground from 85 to 95 per cent passing 200-mesh, depending on mix composition. Some cement chemists believe that raw materials cannot be too he, since finer particles theoretically combine more readily and intimately in the kiln. However, there are certain definite practical objections to extremely fine kiln feed, such as loss of kiln dust, unstable material flow through the kiln and greater dificulty in control of chemical composition due to dust loss, and the disproportion of fines from the more easily ground component. Our

Jan 1, 1947

By John Wulff, Robert Kamm, M. A. Steinberg

IN a previous paper1 from this laboratory, the development of a lead grid method for exploring the distribution of strain and density within metal powder compacts was described, In that work thin lead foil with rectangular holes cut at regular intervals was inserted with the powder fill, the arrangement pressed and then radiographed to permit measurement of the deformation of the lead grid. Densities at each grid element obtained in this manner were plotted as a function of the die dimensions. From such plots it was, found that the densest part of a compact pressed from one side (the top) is at the top outer region and the least dense at the bottom outer region -of the specimen. The axial density maximum occurs more often nearer the bottom than at the top. As a result of such exploratory measurements, it appeared reasonable to conclude that variations in density could be attributed in the main to die wall friction. Thus compacts of large diameter appeared relatively uniform in density except where the die fill had been uneven. Finally it was found that lubrication of the finely finished die cavity contributed more to magnitude and uniformity of density within the compact than interparticle lubrication. Since publication of the above results, it was found that circular-hole grids could be more accurately made and, when deformed within the powder, more readily measured and analyzed. Much of the former work was repeated and new results which include the effect of lubrication, compact height, pressure, speed of pressing and vacuum pressing are reported in the present paper. CIRCULAR-HOLE GRIDS AND ANALYSES OF RADIOGRAPHS Commercial lead sheet 0.008 in. in thickness was used in all the work reported. The 0.0938-in. diam holes were punched in the sheet on 1/8-in, centers to an accuracy of 0.001 in. The finished grid used contained eight or more holes in a row and from I to 24 rows. The lead sheet was clamped between two pairs of brass strips bolted to a small milling machine table which was itself bolted to a bench in correct position near a vise which held the punch. Travel of the lead grid in two dimensions was controlled by two steel lead screws calibrated in 0.001-in. divisions. A stripper plate was used on the punch to avoid distortion of the lead foil. A high carbon-high chromium (chrocar) one-piece steel die was used throughout the work. The die cavity was 1,125 in. in diam and the clearance between plungers and die wall was 0.001 in, The die was also used for pressing from both ends with two long plungers. Finally, lapping of die and plungers gave a surface whose rms value on an Abbot profilometer was 5 micro-in. The die walls and plungers were, except where otherwise noted, lubricated with a stearic acid suspension in benzene. This was put on in multiple layers and allowed to dry before filling the die.

Jan 1, 1948

By Gao K, Wu C, Langley J. M, Waugh B, Farrell K

The dissolution of carbon from carbonaceous materials into molten iron is a key step in various ironmaking operations. For example, in new direct smelting processes, carbon in the form of coal is injected into the molten iron bath, which leads to the dissolution of carbon. In foundry operations, typical recarburizer materials can range from high purity graphite to chars. The dissolution performance of these" materials is critical to the operation and could dictate the efficiency and economics of the process.Dissolution kinetics has been studied by various investigators to explain the influence of melt conditions and carbon type (Sahajwalla et al, 1994; Shigeno, Tokuda, and Ohtani, 1985; Orsten and Oeters; 1988; Ganguly and Reid, 1992; Wright and Taylor, 1993). The influence of melt sulphur content on the dissolution kinetics has been characterised (Sahajwalla et al, 1994), with sulphur retarding the rate of carbon dissolution. The carbon dissolution from different sources is seen to vary significantly even under similar melt conditions. This observation has been explained by investigators on the basis of the varying ash content present in the carbonaceous materials. However, ash content alone cannot account for the variability in dissolution kinetics reported in other studies. One such study is presented below.In a recent study (Sahajwalla et al, 1994), pulverised carbonaceous materials were injected into the molten iron bath and the" dissolution kinetics was established. The dissolution of graphite and glassy carbon, which contain almost 100 per cent carbon, was seen to be significantly different when the molten iron contained sulphur. The influence of melt sulphur content was explained on the basis of different wetting behaviour observed.Importantly, the study showed that even pure carbon sources dissolved at significantly different rates. Hence, their performance could not be explained on the basis of ash content.The key difference between graphite and glassy carbon is in their atomic structure. Graphite is highly ordered whereas glassy carbon has relatively poor ordering in its atomic structure. This study showed that the atomic structure in carbonaceous materials can have a significant effect on their dissolution performance.The focus of the present investigation which stems from the past study described above (Sahajwalla et al, 1994), is to investigate the influence of atomic structure of carbonaceous materials on their dissolution into Fe-C-S melts.The specific objective of this study is to determine the relationship between dissolution kinetics and crystallographic character of carbon sources. In order to accomplish this goal, a range of different materials representing a spectrum of carbon sources have been chosen for the investigation. These materials include brown coal char, two types of Auscarb (A and B), and commercial graphite. Auscarb is prepared by calcining Auscoal at 1000°C. Auscoal is made from chemically cleaned Australian coals (Bowling, Chandler and Waugh, 1990). Part 1 of this paper details the dissolution performance of the materials studied, while Part 11 describes the X-ray diffraction study. The relationship between the atomic structural dimension and dissolution rate constant is discussed in light of the results obtained.

Jan 1, 1996

By Andrew B. Crichton

THE oft repeated statements of the United States Geological Survey and the Bureau of Mines that the coal reserves in the United States are sufficient for 3000 yr have given us all a sense of security hard to disturb. If we discount that statement 66 pct, which to the uninformed seems drastic, we would still have 1000 yr supply, and why should mortal man worry about the difference between 1000 and 3000 yr supply of coal. That is exactly the attitude which many who should be better in- formed have taken toward the problem. A recent brochure "How Strong is America," by the New York Herald Tribune, states there is "no end to coal * * The truth is staggering * * U. S. has enough coal to last 4000 years." What the statement actually means is that the calculator got as high as 4000 yr and decided it was pointless to go on. The President's ERP Committee on Natural Resources and others in high places continue to assure the country of our superabundance of coal. In the past few days the “Wall Street Journal" carried an item from Director Boyd of the Bureau of Mines saying that within a hundred years we would need to triple our annual coal production of 6oo,ooo,ooo tons, making it 1,800,000,000 tons annually, and at that rate we had coal reserves sufficient for 1000 yr. In my humble opinion, we cannot now or ever produce 1,800,000,000 tons annually, be- cause we do not have the reserves in the producing districts to do it. Past reckoning in geology may be, as in the Psalms, a thousand years are as yesterday, but the future must be considered more realistically, and "a thousand years" is really a "heck" of a long time. In most developed areas, mining engineers can count the coal reserves in terms of a hundred years or less, and in some important areas now producing large annual tonnages in much under fifty years. They think when they hear of thousands of years coal reserves, reference is made to distant fields and "the fields are always green far away." The estimates of our coal reserves upon which we are relying were made in 1907 under the direction of Marius R. Campbell, Director of the United States Geological Survey. It is apparent this work was not well done. It was necessarily based on rather meager data because of the little development and prospecting as of that date. There were mistakes in the correlation of seams, and they assumed the continuity of thick seams throughout large coal bearing areas where coal often did not exist. The overstatement of coal bearing areas in the Campbell figures are common. The 1907 survey includes all coals over 14 in. in thickness to a depth of 3000 ft. The total coal reserves according to Campbell were 3157.2 billion tons. The production and mining losses that year would amount to -570 million tons and

Jan 1, 1948

By R. W. Kraft, F. D. George, F. D. Lemkey

From a conszderation of the geometrically possible ways in which an array of lines or linear features in three-dimensional space can depart from a statistically random arrangement, a system was developed to describe and measure different types of anisotrypic arrangements of lines in opaque solids. The technique was applied to a specimen of unidirectionally solidified Al-Cwll, eutectic which contained microstructural imperfections termed lamellar faults. Experimental data obtained from photomicrographs of different facets of a grain of this alloy were then analyzed using this system. It was deduced that the defects, which are essentially linear in natuve, were predominantly parallel to one another but not parallel to the nominal growth direction probably because of some lateral heat flow during solidification. InGOTS of lamellar eutectic alloys which have been unidirectionally solidified under growth conditions which prevent colony formation consist of large elongated grains in which the lamellae are approximately parallel to one another. Deviations from a parallel lamellar structure within a grain are caused by extra lamellae which form occasionally during the solidification process. Three-dimensional analyses of such specimens have revealed that one extra platelet produces two defects, termed 1amellar faults, which can and do grow more or less independently of one another.' The faults are analogous in many ways to edge dislocations in crystals. If a lamella of one phase is considered to be basically two-dimensional, e.g., a surface, then a fault can be defined as the locus of points which delineates an edge (or half the periphery) of one extra lamella as it grows and forms bridges and saddles in an otherwise parallel arrangement oi lamellae, Fig. 1. Were it not for the fact that bridges and saddles occur in the structure, the extra platelet could be defined as the unit defect and the concept of linear faults of positive and negative character around the periphery of the extra platelet would be unnecessary. If an extra platelet is merely inserted (actually it is grown in) between other parallel platelets, the fault line around the periphery of the extra platelet would circumscribe one and only one platelet, namely the extra one, and the faults would lie in a plane parallel to the lamellae. But when an extra platelet merges with adjacent ones and each side grows independently of the other, it is no longer topologically possible to delineate "the extra plate." A fault circuit which either closes on itself or passes out of the specimen still occurs although it no longer will be coplanar with the lamellae. In fact, it is usually not even Planar even if the lamellae are. Under these conditions a concept that a specimen contains N extra platelets of a given size per unit volume would be very artificial and consequently meaningless. Therefore, because of the topological complexity of the alloys, it was deemed necessary to define one-dimensional

Jan 1, 1962

By Wallace Thorneycroft

FOREWORD This paper by Mr. Thorneycroft, Past President of the Institution of Mining Engineers (Great Britain), and chairman of its Subsidence Committee, is a valuable contribution to the assemblage of data on ground movement and subsidence from coal mining. It deals with the effects of advancing longwall, a method almost universal in Europe but as vet little used in the United States outside of Northern Illinois and in a few scattered places elsewhere, in thin coal beds, as in the Lexington field, Missouri, Canon City field, Colorado, and the Birmingham field, Alabama. Although by advancing- longwall, a maximum recovery of coal is obtained and it has to be employed where there are weak shale roofs, especially in deep mines, it is a more expensive method of operation unless the mining conditions are favorable, which is unusual in the United States. On the other hand, there is a growing use in this country of what is essentially retreating long wall, with an approximately straight break-line as described by Rayburn.1 With this method, in which the pillars are substantial, it was shown that the ground movement effect is practically identical with that of advancing longwall, hence the value to mining engineers of the precise data given by Mr. Thorneycroft. His study is unusual in several respects; it deals with conditions connected with ,the mining of a coal bed 20 in. thick, about 490 ft. below the surface, over a mined-out area in a 5-ft. bed about 90 ft. below, which had been mined about 12 years earlier. Both mines passed under Plean House, a building of massive cut stone construction built over a century ago, in which Mr: Thorneycroft resided throughout the under-mining period. Slight cracks opened in the walls as the longwall face approached the house in mining the upper seam, and closed after the face passed and subsidence took place. In the northwest wall, the last wall the face passed under, some stone blocks were broken, which was attributed to a strike period when the advance of face came nearly to a standstill. Other than this there was no perceptible permanent damage, as indicated by the photographs taken by the undersigned in 1923 (Figs. 1 and 2). Mr. Thorneycroft used a novel means of studying the successive positions of the corners of the building: he hung plumb bobs from the eaves and traced the paths of the respective plumb bob points on level surfaces. The angles made by the house walls with the plumb bob lines gave measures of the slope of the subsidence wave at the respective points. A slight upward wave was indicated by the surveys ahead of the subsidence, which would appear to be due to a tilting over of the massive sandstone stratum in the roof measures near the surface, the edge of the subsidence area acting as the fulcrum.

Jan 1, 1931

By J. T. Griffith, Walter S. Lutz

The rainfall during the last six months of 1942 in Scranton and vicinity was 24.06 in.—only 4.35 in more than the average for this district in any equivalent period—yet that rain forced The Hudson Coal Co. to handle 38.8 tons of water for every ton of coal mined. It filled to overflowing, or to a higher level, 14 flooded mine areas holding, in total volume, 8.3 billion gallons of water in storage; brought about a pumpage of 11.4 billion gallons in the months of December 1942 and January, February and March 1943 Half of the total energy generated or purchased (44,322,980 kw-hr.) was required to keep the mines sufficiently free of water to permit the mining of coal during a period of much needed output. Fig. I shows the extent to which this flooding of exhausted and idle properties has proceeded, and is known to be proceeding further, at the upper end of the Northern field. This rainfall (24.06 in.) means 80,559,444 tons of water on approximately 46.2 sq. miles of area tributary to the pumping plants. During the high rainfall period, 72,687,661 tons went underground, being pumped at the stations in this territory. This means an inflow of 88per cent, a runoff of 12 per cent compared with 64 per cent during normal conditions in a normal region. During this period no floods, either surface or underground, had to be dealt with, although the streams and rivers were high on several occasions. In general, the rainfall for the first three months of 1943 can be described as heavy but not abnormal.9 These figures not only indicate the large volumes of water to be handled but, by their very size, they also emphasize that the mine-water problem is not to be conquered, at least from an economical standpoint, by merely buying and installing more and more pumps. Into the solving of this water problem must go much planning, adhering to the principle of keeping the highest percentage of water at the highest possible level for the longest period of time; the storage of huge volumes in sumps capable of being reduced after the flood inflow and at rates less than the inflow; plans for prevention of surface inflow; provision of the minimum of pumping capacity arranged to as nearly 100 per cent pumping load factor as is possible; intercompany joint pumping plants; the conducting of water from one plant to another and from one property to another over distances measured in miles; priorities in power to the threatening situations; and. balance of the necessities and dangers existent in one colliery over another, day by day, when quantities of water are running high. Despite these astronomical figures of water volumes, their disposal to the surface is accomplished reasonably cheaply because, although the annual cost of $575,020 is a large sum, the cost per ton is about 0.1168 cents. Credit for this goes

Jan 1, 1947

By W. J. Parton

The efficient recovery and preparation of small sizes of anthracite called No. 4 Buckwheat (3/32 by 1/32 in.) and No. 5 Buckwheat (1/32 in. by 0), pre-sent a difficult problem to the anthracite operators. In many instances preparation of these sizes, particularly No. 5 anthracite, is extremely inefficient. In many of the older preparation plants, inadequate facilities, and often none at all, were provided as an integral part of the plant for cleaning this fine anthracite. Consequently, much of the material smaller than 3/32 in. was discarded. As the demand for these sizes increased, facilities for preparing the fine sizes were often crowded into existing structures with consequent sacrifices of capacities and efficiency. In addition to this condition, the methods of preparing the finest sizes are being developed in an attempt to improve their efficiencies. Because the demand for small sizes of anthracite still is increasing, many operators would benefit by investigating and improving the cleaning equipment employed on these sizes. The object of this paper is to de- scribe a stationary screen device from which the underflow is restricted by orifices, called a launder screen, along with operating results of several applications of this screen; a settling tank in which two classified products are produced by the installation of a cylindrical partition of smaller diameter than the original tank, one product being recovered from the small tank created by the partition and one product from the space between the inner and outer tank walls; a small pocket classifier or hydraulic trap for use in removing high ash particles from a sludge flow; the classification or segregation of sludge solids occurring in a 16 in. diam pipe flow by analysis of samples procured at different zones; and the operating results of a Fahrenwald sizer, as well as the performance of the concentrating tables handling the classified products. These simple devices were employed at the collieries of the Lehigh Navigation Coal Company Incorporated, Lansford. Pa., in order to improve preparation results on the fine sizes of anthracite. Launder Screens The launder screen is a screening and dewatering device which probably could be used advantageously in the flow diagram of many fine coal plants. Fig 1 shows that it consists of a stationary screen constructed by placing 6 in. high partitions every 6 in. along a launder. Holes are then drilled on 6-in. centers across the bottom of the launder to receive pipe bushings of the desired diameter. The screen cloth is tacked on top of the partitions. When water and solids are fed to the screen, only a small portion of the total water is removed through the orifices in each compartment so that the water in the feed is distributed evenly over the whole length of the screen. The screening action which results is very efficient because the solids are kept in a fluid condition for the full length of the screen. Blinding of the screen can be minimized by the use of screen cloth

Jan 1, 1950

By J. G. Mravec

The SO-called gas-tube method as developed by Hare, Peterson and Soler for determining the type and content of gases in molten steel is particularly adapted for determining the hydrogen content in molten steel. The development of this method has been quite completely discussed in a paper published by these co-workers in 1937.' The purpose of this paper, therefore, is to familiarize with this method those who arc not well acquainted with it. The principle of the gas-tube method is as fol1ows: One end of an evacuated gas tube is sealed with a very thin copper disk. when this end of the tube is immersed in a test spoonful of molten steel, the copper disk melts almost immediately, allowing a portion of the steel to be drawn up into the evacuated space. Any dissolved gases, and those produced by reactions between the components of the steel during freezing, are liberated into the unfilled portion of the sampling tube. A blob of steel that forms on the immersed end of the tube after the sample is drawn up seals up the end, making it airtight again. The evolved gases are subsequently drawn out of the gas tube through a suitable valve and analyzed by means of an Orsat apparatus. The volume of the sampling tube is sufficiently large so that a pressure considerably below atmospheric is produced after the gas evolution. The low pressure is necessary to effect a nearly complete removal of gases from the steel. Construction oF Sampling Tube The sampling gas-tube apparatus is illustrated in Fig. I. Seamless S.A.E. tubing used for section A is first cut into 7-ft. lengths. Through cxperiencc it was found necessary to use two sizes of tubing for this part, "is-in. and 3/4-in., respectively. The smaller tubing is generally used on the hotter or lower carbon steels. If the inside wall of the tube A is coated with grease or oil, it should be burned out before being machined as follows: One end is machined on the inside diameter to a depth of 34 in., so as to form a seat for a 7/8-in copper disk, which is stamped out from 0.005 to 0.007-in. thick copper sheet. The other end is machined on the outside diameter to a diameter of 1 5/16 in., to take a small reservoir or trap member I?. About 4 in. below the same end, a 1/2-in. hole is drilled to receive the 1/2-in. copper tubing leading to the large reservoir C. The inside of tube A is then thoroughly cleaned with a round wire brush, such as is generally used to clean out shotgun barrels. It is very important that the surface be absolutely free of rust or any nonmetallic coating. Next, the thin copper disk is soldered in place with low-melting silver solder.

Jan 1, 1945

By R. S. Pratt

IT is proposed to discuss the materials and furnaces, as well as the melting procedures used in preparing metal for casting billets, bars or cakes for mill fabrication. As far as possible, consideration will be given to the alloys that are commonly made in most brass mills producing strip, tube, rod and wire. RAW MATERIAL The basic raw material for such mills, of course, is copper, which may be obtained in the form of ingots, ingot bars or cathodes. Refinery-cast ingots and ingot bars, which are simply double ingots, are supplied in accordance with A.S.T.M. Specifications as regular electrolytic or fire-refined tough-pitch copper. In the wrought-metal industry, the electrolytic brands have been preferred, although during the war large quantities of fire-refined coppers sufficiently pure to be suitable for all existing requirements were imported from South America. With the exception of Lake copper, most primary fire refined coppers produced in this country have not been considered satisfactory for wrought metal alloys. Cathode copper, which is now widely used instead of ingot copper for alloying, is the electrolytically deposited metal that is normally remelted in the production of electrolytic tough-pitch ingots. It is in the form of rough, flat plates not suitable for processing but quite satisfactory for melting when sheared into suitable sizes. Raw zinc is perhaps the next most important raw material and this is obtained in various grades, depending on the analytical requirements of the alloy to be made. The grades of zinc as outlined in the A.S.T.M. Specifications vary in cost with the amount of impurities-lead, iron, cadmium-and are purchased in such a way as to produce the most economical mixture consistent with the purity required in the alloy. Other virgin metals (lead, tin, aluminum and nickel) are used in varying quantities for making copper alloys. These elements can also be purchased in accordance with the appropriate A.S.T.M. Specifications where available. Also available are a number of special alloys, which are used as deoxidizers or as means of adding relatively small quantities of elements not suitable for adding directly as virgin metal. In this category are such materials as phosphor copper or phosphor tin, silicon copper, manganese copper, cadmium copper, and copper-iron or zinc-iron alloys. Some of these alloys may be purchased on a definite alloy basis or may be produced in the plant itself for its own use. For largely economic reasons, however, a great proportion of the charge to the melting furnaces consists of alloy scrap. In the production of many finished products, as much as 70 or 75 per cent of the metal originally cast may be discarded at various points along the production line. This material eventually finds its way back to the casting shop as mill scrap, overhaul or scalping chips, skeleton scrap, rod turnings, or miscellaneous fabricating scrap. One of the most serious problems in the

Jan 1, 1946

By Irwin Roman

In studying the possibilities of a continuously recording magnetometer for use along the surface of the earth and in an airplane, the Federal Bureau of Mines was led to a study of the theoretical resolving power of magnetic observations, both in the variation of the total magnetic intensity and in each of the two cornponents ordinarily measured. Along with this general study, an attempt was made to determine the relative worth of the total intensity anomaly and the vertical anomaly in geophysical prospecting. Genera1 considerations led to three conclusions: I. In the northern magnetic hemisphere of the earth, the total intensity anomaly is offset to the south of the vertical anomaly, for a positive anomaly (see Appendix, note 1). 2. The resolving power of magnetic observations decreases rapidly with distance from the disturbing body. 3. The horizontal anomaly has little value in prediction of the disturbing body but might be useful in making a choice between bodies suggested by the vertical or total anomaly. Some of the considerations leading to these conclusions were: I. The normal intensity and the anomalous intensity are vectorially additive; hence for a positive anomaly in the northern magnetic hemisphere the total intensity will be a maximum at some point south of the disturbing body and north of the point at which the line to the disturbing body is parallel to the normal earth's field. The anomaly itself has a maximum value directly above the body. For a positive anomaly of constant magnitude, the total intensity will be a maximum when the anomaly vector is parallel to the normal vector. This is shown schematically in Fig. I (not to scale) for a magnetic distribution not restricted to a relatively thin sheet in the magnetic meridian. The normal field has a magnitude 10 with vertical component V0 and horizontal component Ho. Six station points along a traverse across the body from south to north are shown at M to S. At each of these stations the normal magnetic field vector terminates at F. At each of these stations the magnetic field vector due to the disturbing body B terminates at A. The line through the station point and the point A passes through B. The magnitude of the anomaly, indicated by the length of the vector terminating at A, is proportional to some power of the distance from the station point to the disturbing body. If the disturbing body B is a concentrated point of magnetic material, the anomaly is proportional to the inverse second power of the distance. For a uniform magnetic distribution over a line perpendicular to the observed traverse, the anomaly is proportional to the inverse first power of the distance. For surface distributions, the law of intensity is more complicated. At each field point, the total magnetic intensity T is the vector sum of the

Jan 1, 1949

By L. M. Barker, E. G. Lewis

The literature of milling is replete with papers devoted to the subject of ball milling, all of which no doubt have contributed in one way or another to progress in that art. In this paper reference will be made to only a few of the many authors, but the value of the contributions of all who have written on that subject is hereby acknowledged. With a few exceptions the efforts of most students of the subject have been devoted to laboratory investigations of principles. Davis1 was one of the first to report on plant-scale investigations and to present to mill men certain criteria for the guidance of ball-milling practices. His conclusions with respect to the significance of sizing analyses of classifier sands in their relation to the factors of mill speed and ball sizing are particularly referred to. Those conclusions were: 1. If the balls are large or the speed of the mill is high, crowding will appear at the finer sizes in the classifier sands. 2. If the balls are small or the speed is low, crowding will appear at the coarser sizes in the classifier sands. 3. The indications are that best efficiency is obtained when the screen analysis of the sands show a minimum crowding at any size. The last conclusion was qualified by the statement that it had not been proved conclusively. It is surprising, however, that the litera- ture of later years reveals little evidence that millmen have made any serious attempt to apply those criteria. This may, perhaps, be explained by the fact that the average plant-scale ball-milling problem presents so many variables that an investigation of any one of them encounters the problem of control of the others—a problem that often is unsurmountable in a small plant. In a large plant, in which there is a duplication of grinding units in parallel operation, opportunity may be afforded for determining the value of such criteria. Nevertheless, a study of published data from grinding circuits of both large and small plants suggests that either Davis' criteria have been tried and found inapplicable or have never been given consideration. For the purpose of testing Davis' conclusions and with improvement of ball-milling efficiency as the final objective, a test program was initiated in the concentrator of the New Cornelia branch of the Phelps Dodge Corporation early in 1938. The practices of that plant have already been described.2 The fine-grinding units are arranged so that any one of eight units may be operated as a test unit and its results compared with those of the remaining units. All units receive similar feed from a common storage bin, in which the product of the fine-crushing plant is bedded by traveling tripper. The ball mills are all similar in size and operated at the same speeds, and are individually metered. The feeder conveyors of each unit are equipped with Merrick weightometers. The test program as initially outlined proposed investigation of the following:

Jan 1, 1943

By J. B. Pullen

THE United Verde mine is in the north central part of Arizona, on the northeasterly slope of the Black Hills, near the town of Jerome. Ore was first discovered in the district about 1875, and the first mining claims were located by M. A. Ruffner and associates in 1877. In 1888 the late Senator William A. Clark secured control of the property and under his ownership it developed into one of the country's most important copper producers. The property was sold to the Phelps Dodge Corporation in 1935, and since then has been under this Corporation's management. Except for shutdowns caused by general world-market conditions, production has been continuous from 189o to date. GEOLOGY The mineralized zone ranges in area from approximately 12 acres to less than 5 acres. All mineralization occurs in the pre-Cambrian intrusives. The first intrusion, according to geologic records, was the quartz porphyry, which later was intruded by augite diorite. Following the diorite intrusion through a series of intrusions of lesser magnitude, a pipelike mass of massive sulphide, quartz and rock was formed. The diorite forms the hanging wall of this mass, and the porphyry the footwall. Between the footwall and the sulphide a band of chlorite schist was formed, probably owing to metamorphic conditions. Commercial ore occurs along the contact of the schist and sulphide extending into the sulphide, sometimes entirely through the schist, and fingering out into the porphyry. The ore deposit was formed by replacement similar to deposits formed in limestone areas. Faulting and deformation occurred in pre-Cambrian times and again in the Tertiary period, removing approximately 2000 ft. from the top of the main United Verde ore body, forming the United Verde Extension ore body, the only other major ore body in the district. Major development work has shown that the sulphide mass extends to the 3300-ft. level, and diamond drilling below this level has indicated that the mass might extend below the 3700-ft. level, but in a diminished quantity. While many copper minerals exist in the mine, the chief commercial mineral is chalcopyrite. A plan and vertical section of the mineralized zone showing the principal geological features is shown in Figs. I and 2. GENERAL DESCRIPTION AND PHYSICAL CONDITION OF MINE After over 50 years of active production, ore still exists on almost every level of the mine. This condition is due to many causes -mine fires, which broke out as early as 1894, caused the abandonment of certain areas; subsidence and ground movements retarded the cleaning up of others. Furthermore, in the early days there was such an abundance of ore that when mining became difficult it was a simple matter to replace the production by going to a deeper level. Several attempts were made to recover the ore in the fire zone, both by glory-holing

Jan 1, 1941

By J. Peperakis

Coal mine bumps are normally associated with pillar mining under moderate or deep cover. Severe bumps at Sunnyside, however, have not been confined to pillar lines. Many have occurred in virgin development and in localities a long way from active pillar workings. As early as 1918 bumps in virgin territory were attributed to the faulted structure in the Sunnyside area.' Sunnyside is an old property, mined for more than 60 years. Both man-made and natural conditions contribute greatly to bump occurrences, and loss of life and great destruction of workings have resulted from some unusual bumps in recent years. Before Kaiser Steel acquired the Sunnyside Mines in 1950, about 21 million tons of coal had been extracted. Large areas had been worked out and equally large areas first-mined. As demand for coking coal gradually diminished, extensive blocks were shut down, and during the low production period of the 1930's mining on a very limited scale was confined to a section of the property close to the coal tipple. Following purchase of the property Kaiser Steel developed the mines into a fully modernized operation. The worked out and first mined areas covered about four square miles extending five miles along the strike and up to one and a half miles down dip from the outcrop. The sections with favorable mining conditions and clean coal had been largely extracted under the selective mining system imposed by economic conditions and the lack of cleaning plant facilities. Large reserves remained on the fringes of the worked out sections and in areas under deep cover and difficult roof. There were also considerable reserves in first-mined areas. In recent years much of Sunnyside's daily output of 7000 to 8000 tons has been derived from the partially mined localities. Many bumps have occurred when pillars have been pulled within the first-mined areas and on the fringes. The bumps have been associated with retreating pillar lines backing up against old workings and pillar lines in old first-mined areas backing up against highly stressed oversized barrier pillars surrounded by smaller pillars. In some parts of the property the upper Sunnyside seam was extensively worked in the past, and sizable areas of old works overlie the present lower seam workings. Abutment pressures from barrier pillars and pillar remnants are transmitted to the lower seam, contributing to bumps and squeeze conditions. Cover: The terrain is exceedingly rough, and cover varies considerably over relatively short horizontal distances. For example, at No. 1 mine the 2000-ft cover line is ony 2700 ft from the coal outcrop, and at No. 2 mine the cover over the main hoisting slope increases from 1000 to 2000 ft in a horizontal span of 1200 ft. Cover over the present mining sections ranges from several hundred feet in the valleys to 2500 ft under the crest of the divides. A short distance ahead of Sunnyside's deepest development down dip the cover reaches 3000 ft or more.

Jan 1, 1959

By G. S. Scott, G. W. Jones

Economic waste caused by mine fires may become considerable,l5 especially if a fire is allowed to spread or temporarily to get beyond control. It is important, therefore, to act promptly whenever a fire is suspected and to make every action count. To contribute toward its extinction effectively it is necessary to know at the earliest possible moment, and continuously thereafter, what is happening at the seat of the fire, which may be in a remote or inaccessible area, and also to recognize what may be expected as a result of any proposed combative measure. In combatting a fire the mining engineer and the chemist should cooperate, first deciding upon the course of action to be taken and then drawing upon the mining engineer's intimate knowledge of the physical structure of the fire area, the cost and time required to carry out proposed operations, and the probable efficiency with which these operations could be performed, and upon the chemist's knowledge for the results to be expected therefrom. After the initial steps to be taken have been decided, it devolves upon the mining engineer to see that they are carried out and upon the chemist to evaluate the results and to keep the mining engineer continuously informed in regard to a variety of matters—such as the composition and toxic or asphyxiating properties of the atmosphere in which the men may be working, whether the fire is increasing or decreasing in intensity and its probable temperature, the quantity of air leakage and the explosion hazards of the fire area. The purpose of this paper is to discuss what the chemist can and cannot contribute as a member of the fire-fighting personnel. The technical literature contains several hundred papers relating directly and indirectly to mine fires, and this fact alone is testimony to the complex nature of the phenomena involved in the birth, life, and final demise of a fire. A fire may be stopped (I) by preventing oxygen from reaching the coal or (2) by removing the conditions necessary for the rapid reaction of coal with oxygen; i.e., decreasing the temperature. All methods of prevention or control necessarily must be based upon one or the other, or both, of these two principles. The chemistry of a fire may be understood readily if the low-temperature oxidation of coal is considered as a starting point; then oxidation at higher temperatures (or combustion); and with these, the composition of the solid and gaseous products resulting from the oxidation or from the effects of the heat liberated. This discussion refers particularly to Pennsylvania anthracite, although much of the material applies equally well to coals of lower rank. Oxidation of Coal Oxidation of coal fundamentally is confined to the surface; that is, the rate at which oxygen penetrates into the interior of a solid block of coal is so slow that it may be ignored."16 Oxidation starts when a fresh surface is exposed, and the rate is

Jan 1, 1942

By M. A. Steinberg, J. Wulff, R. Kamm

In a previous paper1 from this laboratory, the development of a lead grid method for exploring the distribution of strain and density within metal powder compacts was described. In that work thin lead foil with rectangular holes cut at regular intervals was inserted with the powder fill, the arrangement pressed and then radiographed to permit measurement of the deformation of the lead grid. Densities at each grid element obtained in this manner were plotted as a function of the die dimensions. From such plots it was found that the densest part of a compact pressed from one side (the top) is at the top outer region and the least dense at the bottom outer region of the specimen. The axial density maximum occurs more often nearer the bottom than at the top. As a result of such exploratory measurements, it appeared reasonable to conclude that variations in density could be attributed in the main to die wall friction. Thus compacts of large diameter appeared relatively uniform in density except where the die fill had been uneven. Finally it was found that lubrication of the finely finished die cavity contributed more to magnitude and uniformity of density within the compact than interparticle lubrication. Since publication of the above results, it was found that circular-hole grids could be more accurately made and, when deformed within the powder, more readily measured and analyzed. Much of the former work was repeated and new results which include the effect of lubrication, compact height, pressure, speed of pressing and vacuum pressing are reported in the present paper. Circular-hole Grids and Analyses OF Radiographs Commercial lead sheet 0.008 in. in thickness was used in all the work reported. The 0.0938-in. diam holes were punched in the sheet on 1/8-in. centers to an accuracy of 0.001 in. The finished grid used contained eight or more holes in a row and from I to 24 rows. The lead sheet was clamped between two pairs of brass strips bolted to a small milling machine table which was itself bolted to a bench in correct position near a vise which held the punch. Travel of the lead grid in two dimensions was controlled by two steel lead screws calibrated in 0.001-in. divisions. A stripper plate was used on the punch to avoid distortion of the lead foil. A high carbon-high chromium (chrocar) one-piece steel die was used throughout the work. The die cavity was 1.125 in. in diam and the clearance between plungers and die wall was 0.001 in. The die was also used for pressing from both ends with two long plungers. Finally, lapping of die and plungers gave a surface whose rms value On an Abbot profilometer was 5 micro-in. The die walls and plungers were, except where otherwise noted, lubricated with a stearic acid suspension in benzene. This was put on in multiple layers and allowed to dry before filling the die.

Jan 1, 1949