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By F. G. Rinker

World demand for metallized iron continues to increase, keeping pace with steelmaking requirements for lower-cost hot metal. Virgin iron ore and in-plant oxide fines can be reduced to direct reduced iron (DRI) and hot-briquetted iron (HBI). Maumee Research & Engineering Inc. (MR&E) developed an improved coal-based process for the cost-effective production of either DRI or HBI. Starting with dry compaction, the process combines the latest technology for material preparation, rotary hearth furnace construction and environmental control. Coarse iron ore fines are blended with pulverized nonmetallurgical coal to produce a highly metallized (>90%) product. This paper presents production and installation cost data based on operating and under-construction iron oxide reduction facilities.

Jan 1, 2002

By C. W. Phillips, R. S. Busk

WITH most cast metals the grain size may vary within wide limits, depending upon the conditions at the moment of freezing. These conditions are subject to control in magnesium-base alloys, by proper melting and superheating techniques, enabling production of quite uniform and fine-grained castings, almost independent of section thickness. However, such variations of grain size as do exist, in common with all metals, are reflected in corresponding changes in mechanical properties. It is the intent of this paper to present data giving the relationship between grain size and mechanical properties. Data are also included on the combined effects of grain size and microporosity. In addition, a short discussion of factors influencing the grain size of sand castings is included. MEASUREMENT OF GRAIN SIZE By grain size is meant the statistical distribution of volumes occupied by the individual crystallites in a metal section. Since this is almost always impossible of direct measurement, several methods of estimating and expressing the size have been evolved. Most exact methods make use of the microscope. This involves study of a single plane cut at random through the metal. What is seen in the microscope are polygons, each the cross section of a grain. The diameters or areas of these polygons are then measured and the grain size is expressed as an average diameter, number of grains per unit area, average area per grain, or, rarely, average volume per grain. Rutherford, Aborn, and Bain3 have discussed extensively the errors involved when using a plane section to estimate volume. They clearly point out that the tendency is to underestimate the true grain size because of the nature of the measurement. The A.S.T.M.. recommends that measurement be made by comparison with a standard chart, by means of Jeffries' planimetric method' or Heyn's intercept method. The Society recommends that the notation be in terms of the number of grains per unit area (as is the grain-size number used for steel) or as an average diameter, expressed in inches or millimeters (as for copper and brass). The method of measurement used for this study was comparison with a standard chart, which has been described by P. F. George.', The chart was constructed by photographing at i00 diameters a uniform specimen with an average grain diameter of 0.003 in. The average diameter of this specimen was determined by numerous measurements, using both the Jeffries and Heyn methods. The other grain sizes were then obtained by appropriate enlargement of the master sample. Careful comparison of the chart method with other methods on identical samples indicates that though the differences be-

Jan 1, 1945

By P. G. Chamberlain, E. R. Podnieks, R. E. Thill

Although published data on the physical properties of rock are voluminous, information is often lacking on the environmental conditions under which such data were acquired. Efforts are frequently directed to improve specimen tolerances and preparation procedures; but investigators often overlook giving the rock environment equally careful control during the testing program. Generally, tests are conducted at existing laboratory room conditions, and the prepared rock specimens are frequently, if not most commonly, air-dried for various lengths of time before testing. The standard testing atmosphere recommended by the American Society for Testing & Materials (ASTM)1 is 73.4°F and 50% relative humidity. To maintain a rock at 50% relative humidity for testing purposes, however, requires special equipment and preconditioning procedures, as this study demonstrates. Relatively small variations in the moisture content of a specimen, for example, due to the coolants or lubricants used during specimen preparation, can cause significant variations in the physical properties measured. Such moisture variations may occur in the rock specimen even after it is air-dried, since laboratory humidity conditions commonly fluctuate from day to day. For these reasons, this study emphasizes procedures for preconditioning and maintaining a controlled environment during testing. Although the effects of moisture or temperature on rock properties have previously been studied, their combined effect on rock properties is unknown. In this study, uniaxial compressive strength, Young's and stiffness moduli, and longitudinal wave velocity were determined and compared under controlled conditions of temperature, moisture, and pressure for 11 environments considered to be of particular interest in laboratory, in-situ, and even simulated lunar environments. The following environments were selected for this program: -250°F under dry, 50%

Jan 1, 1972

At Black Lake, Quebec, Lake Asbestos of Quebec Ltd., a wholly-owned subsidiary of American Smelting and Refining Co., has opened a $36-million mining and milling operation which will produce about 100,000 tons of asbestos fiber a year and add 7 pct to the Free World's supply. The orebody once lay beneath Black Lake, and readying the ore deposit for open pit mining meant removing about 37 million cu yd of overburden- 30 million cu yd by dredging and pumping, and 7 million by power shovels and trucks. The enormous amount of material to be dredged, control of drainage during dredging, the need to cope with drainage problems in the area and to provide for flood control to replace the regulating effect of the lake, posed unusual engineering challenges and required tremendous capital investment.

Jan 1, 1959

By Alvis L. Lisenbee, Colin J. Paterson, Mustafa M. Hariri

"The schist-, sediment-, and intrusion-hosted Tertiary Au-Ag hydrothermal deposits of the northern Black Hills, South Dakota are localized and controlled by various types of fractures. This study is an attempt to distinguish between and discuss the possible causes of these fractures in three mines: Richmond Hill, Annie Creek – Foley Ridge (ACFR), and Gilt Edge.Five models based on the structural and tectonic settings of the northern Black Hills are proposed for the fracture origin. In Richmond Hill mine, mineralized fractures within the schist-hosted deposits are in the form of veinlets and stockworks. Larger veins filling open spaces along schistosity planes are also present. Fractures in Richmond Hill mine are related to breccia pipe emplacement and influenced by the Precambrian foliation.Fractures in sediment-hosted (ACFR) and intrusion-hosted (Gilt Edge) deposits are mode I systematic joints. These fractures, and similar ones examined in seven locations outside the mines, are related mostly to the formation of the Lead-Deadwood dome. Fracture trends vary from one area to another depending on their location with respect to the dome. Other factors, such as the Black Hills uplift, pluton emplacement, and Precambrian foliation, may also affect the fracture trends. The proximity of the mines to domes formed by multiple intrusions, the associated fracture patterns and density, and existence of breccias, may be used in exploration for similar deposits in the northern BlackHills or other areas of similar geologic settings."

Jan 1, 1995

By Christopher Haycocks

Because of growing pressure from government and consumers to produce more coal, the mining industry must improve existing techniques and devise more efficient mining methods. Mining of the thicker, shallower and geologically superior coal seams in the western United States is being subjected to increasing pressure from environmentalists; and so in order to meet the demand, the coal industry is exhausting the more accessible surface seams and beginning to mine seam of over 1,000 feet in depth and less than. 30 inches in thickness. The longwall mining system is the natural choice for extracting deep and thin coal seams, as it operates almost independently of the overburden rock stress. Due to the added fact that longwall mining is an inherently more productive method than room and pillar mining, many coal producers are looking to it as a means of increasing productivity without a corresponding cost increase. The consistent rise in application of longwall mining in the United States in the last decade illustrates the excellent potential of this technique for the future. American mining conditions are geologically different from those in Europe where longwall mining has been highly developed and extremely successful. Although numerous successes have been recorded in the U.S. with this mining system, a substantial number of failures have also occurred. These failures are due primarily to roof control problems usually over the face, in the tail entry, or both. Another problem area for American operators contemplating the use of this method has been correct selection of equipment, which should be custom fitted to local mining conditions for optimum productivity. It is essential that correct estimates for chock resistances, support densities, supported and unsupported roof spans, as well as rates of advance and length of working faces be provided to optimize mining conditions. Unless extensive pilot projects are planned engineers must choose their equipment correctly at the outset. A typical investment in longwall equipment will amount to several millions of dollars.

Jan 1, 1979

By B. W. Clyburn

"Coals from the Corporation's Lingan and No. 26 mines are classified as high-volatile A bituminous, with a very low inherent ash and containing a/l essential properties to provide metallurgical coal, with the exception of a high sulphur content. Only through proper cleaning at low specific gravity can the pyritic sulphur fraction of the coal be liberated to an extent where final sulphur meets coking standards. This first section of the paper deals with the operation of a Dutch State Mines heavy-media cyclone circuit to achieve an actual separation in the region of 1.30 specific gravity. A discussion of the parameter of control which must be applied along with an appraisal of the results obtained is included in the paper.IntroductionVery high fluidity and unusually low ash are the outstanding characteristics of coal from Devco's mines in the Harbour seam of the Sydney coal field.However, sulphur content, mainly pyritic, ranges from 1 to 2.5070. To obtain the low and consistent sulphur level required in metallurgical coking coals, preparation at low specific gravity is required .Metallurgical coke is used as the fuel in an iron blast furnace. Its sulphur content largely determines the quantity of slag required to remove it and therefore influences furnace productivity and unit costs.Typical data on the various coal properties of importance in the production of coke are included in Table I. Necessity of Low-Gravity SeparationThe design of any coal preparation plant requires the study of washability properties of the coal to be treated to determine the quantity and quality of coal produced at a given specific gravity along with the difficulty in accomplishing that separation.Their low ash content allows the important washability aspects of Harbour seam coals to be represented by sulphur! specific gravity curves (Fig. I). This particular example indicates that a yield of 54.5% is attained, at a sulphur content of 1.20%, for a perfect separating device operating with a 1.29 S.G. cut point.Material rejected from this separation still contains usable coal, which is recovered as middlings by separating at a higher gravity."

Jan 1, 1980

By J. R. Tickner

"Gold is conventionally analyzed using fire-assay or other destructive chemical methods that impose significant sample preparation, OH&S, and skilled labour requirements. Developed over the past 15 years by the CSIRO, PhotonAssay provides a new approach. A high intensity, high-energy x-ray source is used to activate atoms of gold and other elements in bulk mineral samples. The resulting activity is counted to automatically, accurately and rapidly determine metal grade. Samples weighing 0.5–1.0 kg can be measured in a few minutes with minimal preparation. As the method is matrix-agnostic, traditionally hard-to-assay materials such as highly refractory ores, carbon pulps and copper concentrates require no special treatment. We present the status of commercial PhotonAssay systems being developed by Chrysos Corporation. The first unit with a throughput of up to 80 samples per hour is to be deployed in an Australian assay laboratory in late 2017; smaller systems suitable for providing real-time assay directly at the mine-site will be progressively rolled out from mid-2018. Performance results from extensive testing using both mine-sourced and standard reference materials are presented, demonstrating the sensitivity and accuracy of the method. Implications for replacing conventional gold assay in exploration, mining and processing operations are discussed.INTRODUCTION Accurate methods for assaying ore grade are essential throughout the exploration, mining and mineral processing value chain. The discovery and evaluation of new ore deposits, reserve and resource definition, mine operations planning, ore processing, metal accounting and product quality evaluation all rely on accurate measurement of the valuable metal content of bulk materials. The analysis of gold presents particular challenges, stemming from the very low concentrations at which commercial operations run (often part-per-million (ppm), or equivalently one gram-per-tonne levels or below), and the highly heterogeneous nature of many gold deposits. These challenges mean that any analytical method must have high sensitivity and be capable of measuring large-mass samples. The most widely used quantitative analytical method for gold is fire-assay (Hoffman, Clark and Yeager 1998). An initial ore sample typically weighing several kilograms is dried, crushed and pulverized to allow a small sub-sample (10–50 g is common) to be representatively drawn. The sub-sample is mixed with flux, litharge (lead oxide), a reducing compound and other reagents. The mixture is fired at a temperature of 1000–1200°C until molten. After cooling, the slag is removed and the lead button containing any gold present in the sample is remelted in a porous cupel. The lead is absorbed into the cupel, leaving a gold prill which is either weighed directly, or dissolved in acid and analysed via atomic absorption spectroscopy (AAS) or inductively coupled plasma spectroscopy (ICP or ICP/MS)."

Jan 1, 2017

By Gago P. Oliver

El origen del sostenimiento con pernos para roca se remonta al inicio de la década del 70, donde el uso inicial fue a través de varillas "L"de diámetro de ranura a un extremo introducido a una cuña. El requisito para el desempeño del perno fue contar con un hueco de perforación a la medida y una roca competente para este fin. Posteirormente se realizaron otras pruebas para determinar la capacidad de anclaje de la roca encajonante como de las vetas de l/8" de diámetro con "mariposas" a un extremo como medio de anclamiento. En este sentido, el texto señala que el uso de tubos para el yacimiento de Quiruvilca es favorable, en comparación a los pernos, puesto que mejora el rendimiento y seguridad de la mina, aunque su adquisición sea costosa. Mientras que desde la perspectiva de la mecánica de rocas, el método de reforzamiento es el más efectivo ya que actúa contra la totalidad de la superficie del hueco.

Nov 8, 1984

By Wilfredo Serrano

El presente artículo describe la historia, características, especificaciones, ventajas de la máquina WS Morococha de gran volumen. Dicha máquina fue diseñada en 1966 en el asiento minero de Morococha y es utilizada para concentrar minerales por el método de flotación de forma económica. Entre las principales ventajas tenemos: gran volumen y capacidad por unidad, bajo costo de inversión y mantenimiento, la operación es independiente en cada máquina, su alto radio de concentración permite obtener rougher concentrados limpios directos y que en las paradas no se pierden el volumen del mineral retenido y el punto se descarga en la parte superior. De forma que esta máquina concentradora de minerales por flotación de concepción y desafíos mejorados, desarrolla mayor eficiencia y selectividad.

Mar 1, 1976

By E. G. Woodruff

(Butte Meeting, August, 1913.) FEW authors of treatises and papers on engineering subjects have . given adequate attention to topographic maps.. The statement applies especially to mining engineering in all branches. Even those who have discussed such maps have treated the subject only in a general way. Therefore it is proposed to outline in this paper a few of the uncommon, yet important relationships which topographic maps and their by-product, structure-contour maps, have to modern methods of mining investigation and development when the facts depicted on such maps are properly interpreted. Topography, as suggested by the etymology of the word, menus a detailed description of particular places. Written descriptions have been found less effective than the pictorial representations, therefore attempts have been made in various ways to picture the surface features of places. Lines and shading have been used, hachures drawn, and, finally, the contour topographic map. To a large extent this style of snap is the result of the demands of engineers. It is designed to meet their needs far more than those of the man untrained in engineering. In fact, the average man obtains a better idea of the topography of an area from the hachure system than from contour maps. . Since the maps. have been made chiefly for the use of the engineer, they ought to meet his demands on the one hand, and, on the other, should be used by him to the fullest extent. Such maps, when properly made, accurately portray a portion of the earth's surface and should present the topography better than a personal examination of the area without a map could present it, because the person making such examination views only a limited portion of the surface at a time, and estimates distances only roughly with his eye. Let the engineer consider carefully the fact that in many instances the topographic map is superior to a personal examination of the surface features of a property because the map is based on careful instrumental measurements, whereas the eye gives imperfect data because it can be used merely for the estimation and

Jan 6, 1913

By Robert J. White

Strong's equation for calculating bottom-hole pressure reduction due to gas cutting of drilling mud is corrected, resulting in a simpler equation which is easier to use. Use of the equation is illustrated by examples. It is shown that reduction of weight of a 10 lb/gal mud by 50 per cent at the surface due to gas cutting produces only n 3 per cent decrease in hydrostatic head at the bottom of a 5,000-ft well. It is also shown that at the surface 5 volumes of air per volume of mud are required to reduce the apparent mud weight from 8.5 to 6.5 lh/gal in a 3,000-ft well. Illustrative graphs are included to show the effect of various degrees of gas cutting on apparent mud weight at depths of 2,000, .5,000 and 10,000 ft and to show variation of apparent mud weight, with depth for a 100 lb/cu ft (13.37 lb/gal) mud gas cut to 90 lb/cu ft (12.03 lb/gal) at the surface. INTRODUCTION In 1938, Strong' published the following equation for calculating reduction in bottom-hole pressure due to gas cutting of drilling mud h = 1/D [p + pn/100 In p + p(1-n/100)/p(1- n-100)] where: h = depth in feet P = pressure in atmospheres at depth h due to mud column only D = hydrostatic pressure in atmospheres of a col-umn of uncut mud 1 ft high p = back pressure at well-head (in atmospheres) n/100 = fraction by volume of gas in mud at well-head at back pres-sure p This equation was recently found to contain an error due to an incor-rect expression for variation of vol-ume fraction of gas with pressure. Values calculated by the original equation are close to the correct values for small percentages of gas for which it was derived but become increasingly erroneous as the gas content increases and are completely unusable at the very high gas con-tents employed in aerated mud. DERIVATION The nomenclature used is the same as in Strong's paper. The pres-sure dP exerted by a lamina of thickness dh at a depth h ft is ex-pressed by Eq. 1 where x = volume fraction of gas at depth h dP = D (l - x)dh . . . (1) When we express x in terms of the wellhead per cent gas n, the well-head pressure p and the pressure P at depth h, the result is Eq. 2. x = (np/p+p)/(100-n) + (np/P + p) When this is substituted in Eq. 1 and rearranged to separate the var-iables the result is dh = 1/D [dp + np/100-n dp/(P + p]. Integration between appropriate lim-its results in or if the wellhead is open so that p = 1 atm and we rearrange Eq. 4, the result is loss in head = hD — P = n/100-n In (P + 1) ...(5) Eq. 5 can be solved easily by successive approximation using hD as a first approximation for P on the right-hand side and calculating a first value for loss in head. This can then be used to get a new value of P using the left-hand side of the equation. Values converge very rap-idly even for quite large values of per cent by volume gas n. EXAMPLES The effect on bottom-hole pressure of 50 per cent gas cutting of the mud in a 5,000-ft well using 10 Ib/ gal mud will be calculated, we have: h = 5,000 ft 10 X 7.48 144 X 14.7 = 0.0353 atm/ft hD = 176.5 atm n = 50 Substituting these values in Eq. 5 and using hD as a first approxima-tion of P on the right, we have, approximate loss in head = 50/50 In (176.5 + 1) = 2.30 log 177.5 = 5.17 atm. We calculate a new value of P as follows:

Jan 1, 1958

By Z. T. Bieniawski

In the past 15 years, rock mass classifications have experienced unprecedented development and have emerged as powerful practical methods for assessing ground conditions and tunnel support. Such methods as the RMR System (Geomechanics Classification) and the Q-System have provided many advantages over the classical Terzaghi rock-load method introduced in 1946. These new classifications have found wide applications all over the world but some doubts have been expressed and shortcomings pointed out. The author, who developed the RMR System, takes a penetrating look into the promises and the pitfalls of modern rock mass classifications. He dispels some misconceptions and shows how rock mass classifications can serve as effective and dependable design aids in tunneling.

Jan 1, 1987

By Mathias Jern, Ruilin Yang

A blast vibration model has been developed that simulates effects of almost all blast design parameters on the blast vibration peak particle velocity (PPV) from a production blast with multiple blastholes at a given point of interest in the rock. The velocity of detonation (VoD), initiation points, explosive types (density, VoD, energy), charge decoupling, and decking in each blasthole are simulated along with other blast design parameters. The information from signature hole blast vibration and the blast design are the input to the model. The model can be used to assess blast damage in proximity of blastholes as well as far field vibration PPV.

Jan 1, 2019

By W. S. Versluis, G. W. R. Kuzyk

Atomic Energy of Canada, Ltd. (AECL) is constructing an Underground Research Laboratory (URL), a major component of-the Canadian Nuclear Fuel Waste Management Program. In July 1988. AECL and the US Department of Energy commenced a cooperative research program that included extending the URL shaft by 190 m (623 ft) to the 443-m (1,453 ft) depth. The shaft extension, 4.6-m- diam (15-ft-diam), was excavated by full-face blasting using a burn cut. A Galloway stage accommodated the research activities and the sinking contractor's requirements. Blast designs incorporated controlled blasting techniques to minimize blast-induced damage on the shaft walls. The full-face method has been very effective, with 3.5-m (11.5-ft) rounds being advanced.

Jan 1, 1990

By R. H. Ring

The Automatic Extraction System (AES) was designed and built under a U.S. Department of Energy Contract by the National Mine Service Company in cooperation with the Lee Engineering Division of Consolidation Coal Company. The Pennsylvania State University, Mining Engineering Section, was contracted as an independent auditor of the AES performance. The AES had a rotating cutter drum like a continuous mining machine; but two hydraulically powered roof supports, four roof-bolting drills, a dust scrubber, a ventilation fan and ducting in the roof supports, semi-automatic control, an extensible chain conveyor, and other innovations led to the distinction "extraction system" as opposed to "miner-bolter" or "continuous mining" machine. After design and assembly, the AES was field tested experimentally in an underground coal mine. The in-mine performance audit was preceded by a computer simulation of the potential production for a later comparison with the actual tonnage mined by the AES. This paper describes the results of an in-mine trial and assesses the production potential of the AES.

Jan 1, 1980

By E. R. Stover, J. Wulff

A tentative 870°C isothermal section, the solidus equilibria, and the solubility of Tic and graphite in the nickel solid solution have been determined with arc-cast specimens. Each of the Ni-Ti intermetallics forms two-phase fields with Tic; only Ti,Ni has a measurable solubility for carbon. A quasi-binary eutectic exists between Tic: and a titanium-rich nickel solid solution; ternary eutectics occur on either side, with TiNis or with graphite. Graphite is in equilibrium with a nickel solid solution containing a ratio of Ti/C which increases below the solidus. DURING the past decade, titanium-carbide cermets containing nickel or cobalt alloy binders have been widely studied for high-temperature and wear-resistant applications. Recent work has emphasized control of the carbide grain size and distribution2's and the thermal exparision match between the phases/ The present study was initiated to determine the phase compositions during the sintering or infiltration process, and during subsequent heating. Although the effects of additives and impurities in commercial compositions are important, it was necessary to consider first the pure ternary, Ni-Ti-C. Emphasis was placed on the nickel corner, since the nickel solid solution is the continuous phase in compositions having highest strength. Early metallographic work by Zarubin and Molkov- showed that solid solubility at either end of the section Ni-Tic is small. Edwards and Raine reported that nickel dissolves between 3 and 5 wt pct Tic at 1250° C, on the basis of microstructures of annealed, vacuum-melted powder mixtures. Although the solubility of nickel in Tic has not been measured, Tic formed by the menstruum process in liuid-iron solution contains as little as 0.06 wt pct Fe. BINARY SYSTEMS The most recent determinations of the three binary systems are summarized in Fig. 1. Titanium carbide (6) is predominant among the binary phases, and it melts above the boiling point of nickel. The Ni-C binary is based on separate liquid solubility' ) and solid solubilitylo measurements. The eutectic temperature, 1328OC, is taken from the solidus determination described below. However thermal analysis data by Morrogh and williams also show eutectic arrests in this vicinity in the absence of supercooling. The Ni-Ti binary is taken from the work of Poole

Jan 1, 1960

By C. E. Jordan

The Bureau of Mines is conducting research into the pneumatic recovery of coarse mica and has used this method to produce mica concentrates that contain more than 90 percent mica. This research is being carried out pursuant to the Bureau's objective to develop technology that will help maintain an adequate supply of minerals and metals to meet national economic and strategic needs. Researchers used a Bureau-designed system of crushers, screens, and zig-zag air classifiers to concentrate coarse liberated mica particles from mica-bearing materials. This pneumatic system was used to concentrate four mica ores from Arizona, North Carolina, and South Dakota and three waste tailings from Alabama, Georgia, and South Dakota. Using these samples, it was demonstrated that plus 65-mesh size mica can be effectively recovered by the pneumatic method. Not only were the concentrates high in mica content; it was also demonstrated that this method can be used to recover up to 78 percent of the mica that was originally contained in the samples. Because it is a dry concentration method, the pneumatic beneficiation technique may be advantageous in areas where water resources are limited.

Jan 1, 1980

By T. M. Chance

The first bituminous coal cleaning-plant to use the sand-flotation process1 was placed in operation on Oct. 1, 1925, at the tipple of the East Broad Top Railroad & Coal Co., at Mt. Union, Pa. The general principles governing the design of plants of this type and typical operating results of anthracite sand-flotation units have been previously outlined.2 I will, therefore, confine myself to a description of this particular plant and to some of the elements of design it incorporates. The Original Mt. Union Tipple The plant was installed to transfer coal.from steel narrow-gage cars of 60,000 and 70,000 lb. capacity to standard-gage equipment. The coal was hauled about 30 miles from the East Broad Top field. This coa1, mined by the Rockhill Coal & Iron Co. and other operators, is a low-volatile slow-coking coal with ash of high-fusion point. It is normally a high-grade coal that does not contain an excessive amount of free impurities, but it requires careful handling to secure a fair domestic yield, and, naturally, the need for narrow-gage transshipment increases the difficulty of the problem. The narrow-gage cars were discharged into an 8-car track hopper equipped with two feeders passing the coal to two 42-in. apron intake conveyors. Twin units were installed throughout the original plant, two lump-and-egg shaking screens, two lump picking-tables, two egg picking-tables and two stoker conveyors. The stoker coal was loaded without cleaning. The loading tracks and booms were in duplicate and, with the layout as originally installed, either unit could be run independently, each unit shipping a different class of coal or both units making the same product.

Jan 1, 1927

LIST OF THE MEETINGS OF THE INSTITUTE AND THEIR LOCALITIES FROM ITS ORGANIZATION TO OCTOBER, 1904 Transactions Number Place Date Vol Page I Wilkes-Barre, Pa * May, 1871 I 3 II Bethlehem, Pa August, 1871 I 10 III Troy, N Y November, 1871 I 13 IV Philadelphia, Pa February, 1872 I 17 V New York, N Y * May, 1872 I 20 VI Pittsburg, Pa October, 1872 I 25 VII Boston, Mass February, 1873 I 28 VIII Philadelphia, Pa * May, 1873 II 3 IX Easton, Pa October, 1873 II 7 X New York, N Y February, 1874 II 11 XI St Louis, Mo * May, 1874 III 3 XII Hazleton, Pa October, 1874 III 8 XIII New Haven, Conn February, 1875 III 15 XIV Dover, N J * May, 1875 IV 3 XV Cleveland, 0 October, 1875 IV 9 XVI Washington, D C February, 1876 IV 8 XVII Philadelphia, Pat June, 1876, V XVIII Philadelphia, Pa October, 1876 V 9 XIX New York, N Y February, 1877 v 27 XX Wilkes-Barre, Pa * May, 1877 vi 3 XXI Amenia, N Y October, 1877 vi 10

Jan 1, 1910