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By Thomas E. Marshall, Patrick A. Kinek, Gerald L. Finfinger

The purpose of this investigation was to evaluate the integrity of the existing shutoff valves and actuators used under high-temperature conditions in the Bureau of Mines' horizontal borehole methane drainage pipe- line system and to determine the potential for improvement in the system. In the field tests, both fire-safe ball valves and standard brass ball valves were initially subjected to a 30-min pan fire without thermal protection. Neither valve could maintain line pressure beyond 11 min in a fire that reached temperatures between 500° and 700° C. Based on the test results, it was concluded that the brass ball valve used in the Bureau's methane drainage pipeline system can be retrofitted with an inexpensive "fire bag" to substantially increase its integrity under mine fire conditions.

Jan 1, 1986

By J. L. Carr

On average, one to two fatalities occur every year when a miner is struck or pinned by a continuous mining machine (CMM). Proximity detection systems offer an effective means to prevent these types of injuries and fatalities by sensing the presence of a miner and automatically disabling machine motion. The Mine Safety and Health Administration (MSHA) is expected to finalize a regulation that would require the use of proximity detection systems on all CMMs. Approximately one-fourth of all CMMs in use in the United States are already equipped with such a system. The use of these systems will save lives, not only by automatically preventing accidents but also by acting as a tool for educating miners and reinforcing safe work practices. Researchers at the Office of Mine Safety and Health Research (OMSHR) have conducted a series of field evaluations to quantify the performance of these systems in an underground coal environment. These field evaluations measured the distance at which the warning and stop zones were activated. The technology provided consistent activation of the warning and stop zones under many of the conditions tested. However, some performance inconsistencies were observed which present opportunities to improve the reliability of proximity detection system technology. Some of the notable variables that caused significant changes in performance and should be considered in the design, implementation, and use of proximity detection systems include interaction with the trailing cable, performance changes during tramming, and orientation of the miner-worn component.

Feb 23, 2014

Bendigo Mining operates a 600 000 t/a comminution, gravity, flotation and leaching processing plant for the recovery of gold from their Kangaroo Flat Operation, in the Bendigo goldfields. The free gold and gold bearing sulfi des present are typically coarse grained and hence are highly amenable to gravity recovery. Plant operating data is consistently showing that over 78 per cent of the feed gold is recovered in a circuit, consisting of inline pressure jigs, inline spinners followed by free gold tabling and intensive leaching of tabling tails. The circuit receives coarse crushed product from high pressure grinding rolls and the grinding mill discharge, which allows material to be re-circulated through the gravity process until the particles are small enough to enter the flotation circuit. A plant audit was completed on 26 November 2009 designed to give an indication of the overall operational characteristics of the gravity circuit including single pass efficiencies of associated equipment. The techniques, results and challenges of such a survey are discussed.

Aug 1, 2010

By A. Meléndez

El presente trabajo señala el proceso del entrenamiento para Ingenieros de minas en la corporación minera Cerro Verde a través del plan de perfeccionamiento profesional elaborado por la misma empresa privada. El objetivo del plan es darle al profesional egresado, la experiencia práctica adecuada que complemente la formación académica. Asimismo se busca la eficiencia de su profesión, desarrollar la competitividad y el estímulo de superación del egresado dentro de la empresa minera.

Nov 20, 1972

We live in a world in crisis. One where the tides of change run ever faster. Where the only thing we can be sure of is change itself. The challenge we face, therefore, is not how to resist the forces that move us nor even to try to predict where they will lead us, but to harness their energy and use their momentum to find new and innovative solutions to problems that can seem intractable. Indeed, crisis is often the catalyst which energises a business to change its technology and to form a partnership with a company which specialises in pioneering new technology.

Jan 1, 1993

By E Broers, J Goh, S Davies

With recovery being important in optimising the business performance of a laterite and sulfide plant and with water being a precious resource in Australia, there has been increased pressure for companies to improve their methods of improving liquor recovery and improving water recovery in both laterite and sulfide plants. This has resulted in new control technologies such as the Manta Cube being applied on countercurrent decantation (CCD) and thickening circuits which is necessary to improve stability and optimise the liquor and water recovery in these processing plants. This paper describes the benefit of applying these new control techniques on CCD and thickening circuits in some of the mineral processing plants in Australia.

Jan 1, 2009

By P. Clinton Milne

Uranium mines in Washington State include the producing Midnite and Sherwood mines and several mined-out deposits, the most productive of which was the Daybreak mine. The Midnite mine is generally classified as a hydrothermal deposit consisting of tabular orebodies in Precambrian pelitic and calcareous metasedimentary host rocks. The host rocks occur as roof pendants in a Cretaceous porphyritic quartz monzonite intrusive; the quartz monzonite averages 12 ppm U308 The uranium deposit appears to have undergone a complex genesis, involving a series of depositional and enrichment stages. The Sherwood uranium deposit, approximately 4 miles south of the Midnite mine, occurs near the bottom of a basal Late Cretaceous to Paleocene arkosic conglomerate. The conglomerate lies nonconformably on the porphyritic quartz monzonite that is associated with the Midnite deposit. The Sherwood and Midnite uranium deposits may have had a similar genetic history linked by a common uraniferous pluton source, ground-water action, and/or leaching of uranium-rich ash flows from now-eroded, overlying, Eocene volcanics. The Daybreak deposit, in northeastern Spokane County, consists entirely of secondary uranium mineralization (metaautunite) in a quartz monzonite pluton. Mineralization is confined to a low-angle fault zone and is most highly concentrated at the intersection of primary and secondary shears. Both hydrothermal and ground-water origins are feasible. Besides the Midnite-Sherwood area, 14 other areas in the western Okanogan Highlands physiographic province and Cascades region have favourable characteristics for uranium potential. Criteria include the existence of uranium occurrences, Tertiary basin sediments with intercalated volcanic and volcaniclastic rocks, favourable geochemical data, alkaline plutons associated with uranium occurrences or borders of Tertiary basins, and structural controls.

Jan 1, 1979

By H. A. Saller, F. A. Rough

Studies of the V-U system have been made to determine the constitutional diagram. The diagram is fairly simple, since no intermediate phases are formed. Additions of vanadium lower the uranium melting point and allotropic transformation temperatures. The maximum solid solubility of vanadium in uranium is 12 atomic pct and the maximum solid solubility of uranium in vanadium is about 4 atomic pct. WITH the advent of high purity iodide vanadium,' it became of interest to study the alloying characteristics of vanadium and uranium. Alloys were prepared covering the entire system and were studied using a combination of metallographic examination and thermal and X-ray analyses. As a consequence of this work, the V-U constitutional diagram was evolved. In order to prepare suitable alloys for equilibrium-phase studies, it is desirable to have the purest possible starting materials. For this investigation, calcium-reduced vanadium, de Boer process crystal-bar vanadium, and good Mallinkrodt Chemical Co. uranium were used. The first work was done using the best vanadium then available, that being calcium-reduced vanadium produced at Knolls Atomic Power Laboratory. Subsequently, the important features of the diagram were checked using alloys made with crystal-bar vanadium produced at Battelle. These checks are in agreement with the earlier work. Analyses of the starting materials are contained in Table I. Only minor amounts of any impurities were present in either the vanadium crystal bar or uranium. Nevertheless, care was exercised in selecting the uranium for the alloys. The procedure followed was to section the stock into suitable pieces for melting, and macroetch each piece. Only those pieces which appeared to be clean by macroetch examination were used for alloying. Experimental Procedure Melting: Alloys with 0 to 50 atomic pct V were induction melted in beryllia (BeO) crucibles under vacuum of 10 microns pressure or better. A number of remelts was frequently required to insure uniformity of the individual ingots. Although no visible reaction between the alloy and crucible was found in any of these operations, the alloys were checked by spectroscopic analysis. Only trace amounts of beryllium were detected in the alloys. Alloys containing 50 to 100 atomic pct V were arc melted in a water-cooled copper crucible under an atmosphere of helium. A tungsten electrode was used. Ingots of 10 to 15 g were produced readily by turning and remelting a few times. Such small ingots were necessary because high purity vanadium was not available in appreciable quantities at this time. After melting, the ingots were examined metal-lographically for segregation and were consolidated by remelting before fabrication was attempted. Spectroscopic analyses showed that only small amounts of tungsten were picked up during the arc-melting process. Fabrication of Alloys: Alloys containing up to 20 atomic pct V could be fabricated by rolling in air at 600°C with 5 min between passes. After each pass through the rolling mill the alloys were returned to

Jan 1, 1954

Prof. H. S. Munroe, New York City : In his reference to cop per-dressing at Lake Superior, p. 325, Mr. Douglas says that " the .. concentration .. . has been carried out with greatest financial economy,... I do not say with the greatest economy of min eral." And again, on p. 344, referring to the same, " it would not pay to reduce the quantity treated, and increase hand-labor, in order to save a few hundredths of one per cent. of copper." Mr. Douglas probably has in mind the fact that the copper lost is all included copper, locked op in grains of sand. To save this, the rock must be crushed finer. If the mesh of the stamp-screens be reduced, the stamps will crush less; and thus we shall have to reduce the quantity treated in the mill per day, while the labor re mains the same per day, or is increased per ton, as he says. Submitting to the loss of copper, however, is not the only solution of the difficulty. A large quantity of rich sand is now collected by the jigs, and it would be easy to save so much of this material that the escaping sands would be quite poor. What is wanted at Lake Superior, and at many dressing-works in this country, is an efficient crushing-machine to treat the sands thus collected. Abroad, we have the Heberle machine and the Schranz mill for this purpose, the latter being the more successful. The Heberle grinder was tried at the Calumet and Hecla mill for crush ing these rich sands, with what success I do not know. I should like to ask Mr. Douglas if we have no American crushing-machine that will reduce jig-sands to a fine state of division

Jan 1, 1894

By N. E. Chaouai, M. Lavigne

"Indicator kriging is one of the non-parametric geostatistical techniques, ideal for estimating the reserves of irregular mineralizations (e.g. gold, uranium, platinum, diamonds, etc.). Such mineralizations are characterized by variograms which are very sensitive to extreme values (outliers) and very difficult to model.The classical approach to this problem in the industry is to trim off these high grades using arbitrary empirical upper limits established either from practice or subjectively. Contrary to this simplistic approach, indicator kriging discretizes the histogram of the grades (usually log normal) in several classes and carries out interpolation separately for each class. This paper gives an overview of the differences between ordinary and indicator kriging along with the steps required to carry out indicator kriging. The paper concludes with a case study involving a gold deposit estimation.IntroductionParametric geostatistics (ordinary kriging) has given us satisfactory results over the years in estimating the reserves of well behaved deposits with coefficients of variation close to 1 (e.g. sedimentary and porphyry type deposits). However, numerous problems have been encountered in estimating irregular mineralizations (e.g. small vein deposits of precious metals) with high coefficients of variation in the range of 2 to 5. Such deposits are characterized by a very small proportion of extreme values (outliers) which create distortions on the variograms (pure nugget effect) and render them useless(1,2).The solution often followed by the mining industry today is to trim off these values based on empirical limits. These limits are derived either by comparison with production results (or bulk samples) or by geological interpretations. Needless to say that such manipulations can lead to serious errors in estimating ore reserves (usually underestimation of total metal quantity), especially when production results or bulk samples are absent.On the other hand, the solution followed by the geostatisticians in the past was to smooth out the data using transforms (e.g. normal-score or natural logarithms). Such transforms are nonlinear and they required non-linear estimation techniques such as disjunctive kriging (DK) or multivariate gaussian kriging (MG) which require normal related hypotheses, not necessarily true in geology.Nonparametric geostatistics (indicator and probability kriging) have been -developed in order to handle these problems related to highly skewed distributions."

Jan 1, 1990

By A. D. Hodges

In May, 1850, the first gold from the Comstock lode was discovered in the sands near the Carson river. It had been washed down by natural agencies, through Gold Caflon, from the decomposed rock of the ledge above. Within a few weeks, the precious metal was traced more than three miles in shallow deposits along the bed of the canon, from the river to the Devil's Gate-—a narrow, rocky cleft, notable to the teamsters of a later day chiefly as the place where " Uncle Jimmy Fair" collected high tolls, and could not reduce the rates, because " this was all he had for the education of his poor children." From 1850 to 1857, a fluctuating band of miners, whose number varied from twenty to two hundred, washed the Gold Canon placers with rockers and long-toms, making four to five dollars a day each until the last-named year, when the deposits were in great measure exhausted. The chief settlement of these miners was at Johntown. During only a third of the year was the water-supply sufficient for their work. The remainder of the time they utilized in desultory prospecting of the hills and in occasional forays for supplies among the Mormons in Carson valley. In 1857, the Grosch brothers discovered the Comstock lode at some point, now unknown, in Gold Hill, but their discovery perished with them. Early in 1858, the " Pioneer Quartz Company," having located several of the small surface-veins around Silver City (whose connection with the main lode has never been definitely shown), built an arrastra, and worked some of their ore; but the returns were so small that they soon abandoned this business. The following winter, James Finney and his partners found gold in place in

Jan 1, 1891

The care necessary to ensure a reliable estimate of average value from the individual values of a series of bore holes sunk in any alluvial deposit containing coarse gold was well exemplified in checking over the mining returns from a previously bored deposit in Central Otago, N.Z. While the working returns are always liable to vary considerably from bore values in any particular sub-section of an area, still, over any considerable area, the two must check within small limits, if profits are to be calculated and capital raised on the average value given by boring.The present figures have probably more value than those from, dredge returns since accurate measurements are difficult to obtain in irregular water-filled dredge paddocks, whereas in the case under review the paddock walls stood almost perpendicular, and, as the measurements were taken and checked in the empty rectangular excavation, the yardage estimate may be taken as substantially exact.NATURE OF DEPOSITThe area consists of a narrow valley deposit of fluvio-glacial origin, the original deposit having been eroded and the gold reconcentrated by post glacial rivers. The wash is between 80 and 90 ft. deep, and is formed in two distinct layers:(i.) The top layer from surface to 70 ft. or 80 ft. consists of about equal quantities of gravel mixed with glacial drift, and a small amount of clay. The gravel is typical discoid glacial stone of a maximum size of about 4 in. diam. by 1 in. thick, and consists of soft micaceous schist and laminated slate. This upper layer of wash contains no coarse gold whatever, and very little fine; indeed, in mining, the top 30 ft. is discarded without being passed over the boxes.(ii.) The lower layer of wash to bedrock is similar to the upper, except iihat it contains about 10% of larger discoid stone up to 12 in. diam. by It in. thick; 90% of the gold is concentrated in this bottom layer.

Jan 1, 1929

By A. M. Gaudin

Activation of sphalerite with silver ions results from the exchange of two Ag ions for one Zn ion in the sphalerite lattice. The exchange is proportional to the logarithm of time and proceeds until either of the reactants is consumed. Sodium cyanide prevents the activation of sphalerite by controlling the ratio [Zn++]/[Ag+]' in solution. If sphalerite has already been activated with excess silver, cyanide ions can deactivate sphalerite provided oxygen is present. SOLUBLE: silver salts are effective activators for flotation of sphalerite with potassium ethyl xan-thate as collector.* Since sphalerite takes up large quantities of silver ions rapidly,2-4 investigation of this system appeared fruitful for studying the mechanism of sphalerite activation and deactivation. Experimental data presented here comprise studies of the mechanism and rates of activation and deactivation of sphalerite with silver and cyanide ions. A 20-lb sample of sphalerite from Ottawa County, Oklahoma, was handpicked, crushed, screened, hy-drosized to produce a 65/100-mesh fraction, and purified by electrostatic separation. The sphalerite was then leached with benzene and absolute ethyl alcohol to remove possible organic contaminants and with 0.1 M NaCN to remove any copper that might have been picked up during screening. The benzene had been specially purified by grinding with fine sphalerite for 12 hr to remove trace impurities that might have had a specific affinity for sphalerite. Ii'inally, the mineral was washed, dried in flowing nitrogen, and stored in large ground-glass stoppered weighing bottles under a small nitrogen pressure. The specific surface of two samples from the 65/100-mesh sphalerite stock, measured by krypton gas adsorption," was 420 cm2 per g. Compared with spheres of the same size and specific gravity, the prepared mineral had five times as much surface. Results of a wet chemical analysis and of a spectro-graphic study are given in Table I. Manganese, nickel, tin, indium, thallium, cobalt, and magnesium were not detected spectrographically. All solutions were preparid by -adding reagent-grade chemicals to conductivitv water. which had been previously saturated with the desired gases. Nitrogen and oxygen were purified, but no attempt was made to purify the laboratory air. Radioanalysis was carried out with prepurified Ag110 tracer obtained from the Oak Ridge National Laboratory. The complete decay scheme for Ae is complex. Its primary emissions are 0.09 and 0.55 mev ß rays and 1.48 mev y raysB and its half-life is believed to be 270 days. In preparation of solutions of AgNO3, the amount of Ag110 added was negligible compared to that of the inert silver. Abstraction of silver from solution by sphalerite was measured after agitation of 1.5 g sphalerite with 20 ml of solution in 50-ml pyrex flasks. All sample preparations were carried out in a dry box so that the proper atmosphere could be maintained. After the mineral had been agitated in the silver nitrate solution, an aliquot of the solution and a sample of the wet mineral were transferred to separate counting vials and weighed. The solid sample was dried and re-weighed and the activity in both vials measured by means of a scintillation counter. With scintillation counting it is possible to count y radiation from the solid and liquid, a distinct advantage over the older ß-ray counting techdiques.4 All measurements of silver activity were made using a sodium-iodide well-type scintillation counter. The amount of zinc in solution was measured colorimet-rically by the monocolor method of Sandell.7 Activation Studies Exchange Between Zinc Ions and Silver Ions During Activation: The increase of Zn++ in solution during activation with Ag+ is shown in Table 11. These data indicate a replacement of one Zn++ by two Ag+. Analysis of a separate two-month agitation test indicated that the replacement of Zn++ by Ag+ was, as expected, in the molar ratio of 2.0.

Jan 1, 1958

By William E. Brewster

Dating back some five years ago, various foundries made inquiries as to the probable total carbon content in a given specification and grade of pig iron. Up to that time we had no data, and except for a few occasional determinations of total carbon, no attempt had ever been made to set up routine determinations of total carbon. In fact, at that time, a cast analysis was considered sufficient, without examining at further length into possible variations of analysis between ladles of the same cast. In view of the various inquiries as to total carbon, we sampled all piles of all grades of pig iron, making complete analyses. Finding some variation in carbon content, we secured more samples and plotted the carbon content with respect to silicon in each grade of iron except for the higher phosphorus foundry iron, where it was necessary to divide the group into twenty-point phosphorus ranges. This is easily understood, owing to the falling off of carbon content with the increase in phosphorus content, silicon in each case being the same. In addition, these curves clearly demonstrate the lowering of the carbon content as the silicon is increased. In order to insure accuracy and reliable determinations the following method of sampling and analysis was employed: Two pigs selected from each end of the car, each car containing iron from one 50-ton ladle, are drilled with a 5/6-in. drill all the way through from top to bottom in the center of one section of the pig. The drillings are collected and transferred to an 80-mesh sieve; the portion passing through and the portion retained are placed in separate envelopes and labeled accordingly. The amount of sample in each envelope is weighed to obtain the proportionate amount of coarse and fine drillings, then a 1/2 factor weight of sample (1.3636 grams) made up of proportionate

Jan 1, 1936

By C. Zener, John H. Hollomon

It is commonly recognized that a given material may be described as ductile or brittle only with reference to the conditions of test. Thus under the usual test conditions quartz is brittle, but under high pressures it is ductile. Salts that are brittle at room temperature become ductile at elevated temperatures. Pitch, brittle with respect to rapid loads, flows at low rates of loading. Pearlitic steel, ductile under the usual conditions of test, may be embrittled under the proper conditions of combined stresses, temperature, and rate of loading. Earlier Investigations A clear analysis of the conditions necessary for plastic flow and for fracture was first given by LudwikI1 who introduced two types of stress, the "flow stress" and the "stress at fracture." These stresses may refer to tensile or to shear stresses, or, in fact, to any type of stress. Ludwik considered these two stresses to be independent of one another, functions of the conditions of test and of the prior history of the material. For the purpose of analysis. both a Flow stress and a stress at fracture can be envisaged as functions of strain over the entire range from zero to infinite strain. Thus, the material is brittle if the stress at fracturc is less than the flow stress and ductile if the reverse is true. According to Ludwik's analysis, the problem of determining the effects of combined stresses, of temperature and of strain rate upon the behavior of metals reduces to a problem of determining the effect of these variables upon the flow stress and upon the stress at fracture. In the present paper, methods are described for determining the effects of two of these variables, temperature and strain rate, and examples are given of the dependence of tlow stress and stress at fracture, upon these two variables. The effect of combined stresses will be discussed in a succeeding paper. The flow stress may be determined from a load-diameter curve. Load-diameter curves were obtained over a wide range of temperature and over the limited range of strain rates obtainable with the lever-type Riehle machine used. The curves were then converted to an extremely wide range of strain rate by making use of the equivalence of the effects of changes in strain rate and changes in temperature, previously proposed2 and demonstrated3 by the authors. According to this principle of equivalence, strain rate e and temperature T affect the flow stress only through a single parameter-p of the form: p = eeQ/RT 111 In this equation, R is the gas constant (2 cal. per gram mol per deg. C.), and Q is a heat of activation, which varies from material to material, and may depend upon the strain, but not upon the strain rate or the temperature. There is associated with every value of p a flow-stress curve. article of strain rate and of temperature that give the same p have the same flow-stress curve. Therefore, merely by changing the temperature of test, the effect of changing the strain rate may be obtained. Thus, under isothermal conditions, the material will behave the same at

Jan 1, 1944

By S. P. Sandoval, A. E. Clark, S. W. Yopps, J. S. Gomer

The U.S. Bureau of Mines conducted Cu exchange tests for six common clays under simulated in situ leaching conditions. Regression equations were obtained from the data expressing the Cu exchange capacity as a function of Cu concentration, pH, and temperature. Using these equations, an analysis was made of the impact each clay could have on overall Cu recovery. The results suggest that Ca and Na montmorillonite clays could have a major impact on Cu recovery and that attapulgite and illite clays could have a smaller, but still significant, impact. Kaolinite and ripidolite clays pose little threat to loss of Cu from in situ leaching solutions.

Jan 1, 1992

By Thomas M. Plouf

Flotation, as we know it today, is a physiochemical method of concentrating finely ground ores. The process involves chemical treatment of an ore pulp to create conditions favorable for the attachment of certain mineral particles to air bubbles which are carried to the surface of the ore pulp for removal. Frother reagents, combined with air and the agitation mechanism, produce the effective bubbles in a flotation cell. A method of analysis is presented to evaluate the effect of frother reductions and additions on a bank of flotation cells, and to economically determine if and to what extent frother reductions and additions are justified. Comparative evaluations of different frother systems, as well as the amount of frother used in a given flotation machine, cannot be made on the frother cost alone. An indepth study must be made to evaluate any given frother as applied to a particular ore body. The recovery of a given mineral is as dependent on the amount of frother as on the type of frother used. An actual case study follows in which the amount of frother required by two different flotation machines was used to make a decision as to which unit performed more efficiently based on costs. This same method may be used to evaluate new frothers as well as to determine if a larger frother addition is advisable.

Jan 11, 1975

A hidden asset in most exploration and mining activities is data. Common practice involves the spending of large amounts of money to discover, define and prove the viability of a mining project. Some organisations have used operations research analysis and internal costing methods to allocate a value to their data and to compare the costs of data maintenance to other more tangible activities like plant and equipment maintenance. Those organisations, which do not undertake such an analysis often, have little knowledge or understanding of the extent and the worth of their data. Often the organisations lack of knowledge and understanding of the data is indicated by their failure to fully exploit or realise their opportunities. The focus of this paper is to discuss new technologies that optimise the management and security of collected and derived data. The technologies discussed are components of the software systems acQuire and MineSight.

Jan 1, 1999

MELTING in a crucible is the customary method for small things. It is done in two ways: with the blast of bellows and with an air furnace. The one with bellows, which I wish to show you now, is the common method most in use. It melts quickly and is very well known to goldsmiths and to all melters. I could continue without telling you of it, but I shall tell you its method in order to caution you about what you might perhaps not notice. In order to do this, first make a little forge with a pair of bellows that can be worked with one hand or in some other way. A crucible of the size needed is then taken and filled with the material that you wish to melt. Then, on the forge in front of the place where the wind issues from the bellows, kindle as much charcoal as you think will well cover your crucible and put your crucible full of material in the midst of this burning charcoal above the place where the blast strikes the fire, two or three dita distant from the place where the blast issues-or as much more or less as the size of the crucible or the force of the bellows requires. It is allowed

Jan 1, 1942

By I. M. Roberts

The war emergency has affected every phase of industry. The gas and coke-oven companies have sought faithfully to discharge their responsibility in this critical period and have willingly modified their operations where possible to meet the ever increasing demands made upon them. Our company made commitments for blast-%urnace coke for war purposes which necessitated a reduction in coking time in order to produce the amount required to meet the new demands. At the same time there were certain customers of long standing, who had been supplied with foundry and egg-sized coke produced on fairly long coking periods, who also were engaged in war work and whose patronage we desired to retain. Coke produced on short coking rates from the coal mixture we were then using would not have been satisfactory for these customers. The average width of our ovens is 19¾ in., which is a little wider than the more modern ovens of the present day. They were originally designed to give somewhat cool tops in order to produce a rich gas. This necessitates carrying the heats a little on the warm side so that the tops will be fully carbonized. A reduction in coking time, as is generally known, is accompanied by a reduction in the average size of the coke produced. This is due to the greater shrinkage that takes place at the higher temperatures required to carbonize the charges in a shorter time. The net result is that the coke is more cross fractured and longitudinally fractured and breaks up into small sizes on handling. The cokes produced from the better coking coals seem to be more susceptible to fracture at high temperatures than those from poorer coking coals. A coal mixture that produces a beautiful foundry-sized coke at 30 to 36 hr. may yield a product at 18 hr. that would be so badly fractured as to be unsatisfactory for foundry practice. Coking Characteristics OF Coals Our problem was to find some economical innovation and develop a method that would enable us to reduce the coking time and still produce a coke satisfactory to our foundry trade. However, before discussing the results of our present operations, it might be well to present a few observations on the coking characteristics of coals in general. Most coke-oven operators carbonize a blended mixture of coals. It may consist of a single high-volatile coal with a single low-volatile coal, or of a combination of several high-volatile and low-volatile coals. Medium-volatile coals of the Eagle seam are used for the production of high-grade foundry coke. Medium and high-volatile coals are often blended with one or more low-volatile coals. The possibilities are almost infinite, but the combinations are limited by such factors as cost and facilities for handling multicoal mixtures, and by other practical considerations. Every coal has characteristics that make it blend better with a particular coal or coals in preference to others. This property has

Jan 1, 1944