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By C. E. Meissner

The need for a practical method of deciding upon new alloy steel analyses to widen its markets was the problem facing the Chrome Steel Works at the beginning of 1927. In addition to determining the physical properties of the alloys under consideration, it was necessary to test their value in service. Small melts of steel in a crucible, from which test specimens could be made, were insufficient to determine more than physical properties, therefore actual castings and ingots had to be made before the products could be placed in service. Furthermore, to save further experimental work, it was desirable that these heats should be made on a "pilot" basis; that is, actually duplicating regular open-hearth furnace practice. Therefore a one-ton (2000-lb.) acid open-hearth furnace was built. Furnace Design There are no special features in the design of this furnace. Standard acid open-hearth practice is followed, with the reproportioning of the dimensions necessitated by its small size; notably, a considerably increased length in proportion to width in order to get proper flame propagation. After the first heat was made it was found that the check-ers were insufficient. The checker chambers were increased in size and the checker openings were made much smaller. The furnace was started late in March, 1927, and made 65 heats. Experimental work was then stopped unti1, the data accumulated could be worked up and summarized. The run was resumed in November and the furnace is still operating, having made at this writing 60 additional heats. For the last series of runs, there was some rebuilding of both the furnace itself and the checkers. Slag pockets were added. Fig. 1 shows details of construction; Fig. 2, the general appearance. All designing was done by the operating and engineering staff. Construction materials were those available around the plant. The reversing valve is a standard butterfly type taken from an old furnace. The slag pockets and checkers are cleaned by digging down to arched openings which have been bricked up.

By J. Sibakin, R. D. Hindson

In order to devise a practicable soaking pit schedule for use at The Steel Co. of Canada Ltd.'s Hamilton Works, soaking pit heating temperatures, sooking times, pit capacity, and safe maximum mill drafts were correlated with fluctuations in the current or load of the bloom mill driving motor. Other variables such as total delays in the pit, rolling schedules, mill delays, and track times were also investigated. IN order to show an easily applied and accurate means of establishing soaking pit heating temperatures, soaking times, pit capacity, and safe maximum mill drafts, these various factors are correlated herein with fluctuations in the current or load of the bloom mill driving motor. Rolling practices have a considerable influence on the production capacity of a blooming mill. The maximum values of the torque, in particular, are of importance, since even instantaneous current peaks lead to the tripping of the motor by the overload relay and result in loss of mill time. The establishment of safe maximum drafts and accelerations for ingots of different sizes and of a soaking pit practice which would ensure a consistent and satisfactory plasticity of the metal is of considerable importance for increasing the efficiency of both the blooming mill and the soaking pits. The Bloom Mill Dept. of the Hamilton Works, The Steel Co. of Canada Ltd., is equipped with one 44 in. mill driven by a 7000 hp motor with the setting of the overload relay at 22.0 ka. The speed of rotation of the motor is regulated after the Ward-Leonard system. There are three basic speeds of 9.5, 28, and 47 rpm and a further possibility of increasing the speed by weakening the field. This last possibility is hardly ever used during practical operations. The rolling program of the blooming mill is varied, both in the size of the ingots to be handled and in the steel grades. The total tonnage handled by the mill is about 2,000,000 ingot tons per year. At the time of the investigation, the Bloom Mill Dept. was equipped with 22 soaking pits (6 regenerative, 14 bottom-fired, and 2 one-way top-fired pits) with a total bottom area of 2770 sq ft. The pits are fired with a blast furnace-coke oven gas mixture having a calorific value of 155 Btu per cu ft. The foregoing figures show that the production program was such as to impose the necessity of a most efficient usage of the available equipment. For this purpose, the operations of the 44 in. mill and of the soaking pits were investigated, and the results of the investigation were used as a basis for a revised soaking pit schedule and drafting practice. The plasticity of an ingot of a certain chemical composition when being rolled is determined mainly by the following factors: I—the ingot size, both thickness and width; 2—the length of the gas soak; and 3—the surface temperature. The first two factors determine the uniformity of the temperature distribution over the cross-section of an ingot. The third factor introduces the level of the heating of an ingot. The torque produced by an ingot being rolled is determined by the area of the metal displaced, its plasticity, and acceleration values. On the other hand, with shunt motors the torque is determined by the current. This can be assumed to be correct with only a small degree of error for compound motors with a relatively small effect of the series windings as long as the velocity is not regulated by weakening the field. Since the spread is relatively unimportant when compared to the width of an ingot and since it is also reduced several times during rolling by edging passes, the draft alone and not the area of the metal displaced may be taken into consideration with ingots of a similar size. It is therefore possible to determine the main features of the heating and drafting of an ingot by measuring the current and acceleration of the mill motor. After the acceleration has been taken into account, the amount of current will be an indication of how the motor responds to a heating and/or drafting practice and these practices can be adjusted in order to get the desired result. As peak currents are more likely when heavier ingots are rolled, the rolling of plate and slab ingots was investigated. Conditions prevailing when smaller ingots are rolled can be deduced from the results obtained on heavier ingots. All measurements were made when plain carbon grades under 0.15 pct C were rolled. The motor current, the voltage across the armature, and the rpm were recorded simultaneously on synchronized charts, Fig. 1, which moved with the speed of 6 in. per min. Each draft was recorded by a special observer. The rpm curve made it possible to establish the acceleration at any given moment. For purposes of correlation, the maximum current during a pass and the corresponding acceleration were used. The charts made it possible to establish the position of the roller's lever at any given moment as well as the total time of a pass. The slab ingots were divided into three groups (28x35, 28x45, and 27Mx53 in. ingots) and each group was investigated separately. Since they account for most of the current peaks, only flat passes were used for purposes of correlation, a total of 1373 having been investigated.

Jan 1, 1956

By R. G. Agarwal, Ramey Jr. H. J., Al-Hussainy R.

Although it has long been realized that gas recovery from a water-drive gas reservoir may be poor because of high residual saturations under water drive, it appears that only limited infomlation on the subject has been available until recently. This study was performed to show the qiiantitative potential importance of water influx. Results indicate that gas recovery may be very low in some cases: perhaps as low as 45 per cent of the initial gas in place. Gas recovery under water drive appear to depend in an important was on: (I) the prodirction rate and manner of production; (2) the residual gas saturation; (3) aquifer propertie.); and (4) the volumetric displacement effciency of water invading the gas reservoir. The manner of estimating water-drive gas reservoir recovery can vary considerably. Examples are: the steady-state tnethorl. the Hurst modified steady-state method, and various unsteady-state methods such ac. those of van Ever-dingen-Hurst, Hurst, and Carter-Tracy. The Carter-Tracy water influx expression was used in this study. In certain cases, it appears that gas recovery can be increased significantly by controlling the production rate and manner of production. For this reason, the potential importance of water influx in particular gas reservoirs should he investigated early to permit adequtrtr planning lo optirtize the pay reserves. INTRODUCTION In recent years, the economic importance of natural gas production has become increasingly apparent. This has been evidenced by more intensive exploration efforts aimed at gas production, and exploitation of both deep, as well as low-permeability gas reservoirs. Technical developments such as deep-penetration fracturing have made development of such formations economically feasible. Unfortunately, water influx has forced abandonment of a number of gas reservoirs at extraordinarily high pressures. Although reservoir engineering methods for estimating water influx have long been available, it appears that application of these methods to the water-drive gas reservoir has been sporadic.'a Available methods for estimating water influx which can be applied to the water-drive gas reservoir problem include the steady-state method,1 the Hurst modified steady-state method and various unsteady-state methods such as those of van Everdingen-Hurst. Hurst, and Carter-Tracy. Interesting applications of these solu- tions to gas reservoir and the aquifer gas-storage problems have appeared recently.3,12,14 The experimental study of residual gas saturations under water drive by Geffen et al. in 1952 indicated that residual gas saturations could be extremely high. A value of 35 per cent of pore volume is often used in field practice when specific information is not available. The study of Geffen et al. showed that residual gas saturation might be much higher in some cases. Naar and Henderson concluded that the residual non-wetting phase saturation under imbibition should be about half of the initial non-wetting phase saturation. The Naar and Henderson result that residual gas saturation under water influx should be about half the original gas saturation is recommended as an estimate if laboratory measurements are not available. Thus, it is clear that a considerable portion of the initial gas in place might be trapped in a water-drive gas reservoir as residual gas at high pressure. A full water-drive would result in loss of residual gas trapped at initial reser.voir pressure. Consideration of transient aquifer behavior leads to the conclusion that high-rate production of water-drive gas reservoirs could result in improved gas recovery by reduction of the abandonment pressure. However, there appears to be little quantitative information on this possibility. One of the few advantages of water-drive gas production appears to be improved deliverability through water-drive support of the reservoir pressure. There may also be an advantage in higher condensate recovery caused by pressure maintenance for gas-condensate water-drive reservoirs. In view of the preceding, this study was made to assess the potential importance of water-drive in gas reservoir engineering. The Carter-Tracy approximate water-influx expression was used because this equation offers some advantages in hand-calculation which do not appear to have been generally recognized.' However, calculations were performed in the main with a high-speed digital computer to permit evaluation of the effect of water-drive under a large variety of conditions. CALCULATION METHOD Water-drive gas reservoir performance can be estimated in a manner completely analogous to oil reservoir calculations: a materials balance is written for the reservoir, and a water influx equation is written for the aquifer. Siniltaneous solution provides the cun~ulative water influx and the reservoir pressure. When reservoir performance data (gas produced and reservoir pressures) are available, it is usually possible to match performance data to determine the initial gas in place and the water influx parameters —

Jan 1, 1966

By Michael Tenenbaum

THE importance of semikilled steel as a high tonnage grade has long been recognized. The increasing severity of the applications for which semikilled steel is used makes it desirable to obtain further information regarding the features of open hearth practice and ingot structure that can affect steel performance in subsequent rolling and fabricating operations. Accordingly, an investigation is being carried out studying the structure of various types of semikilled steel ingots. This paper reviews some of the observations made and information gathered in this investigation. By definition, semikilled steel may be considered to include all nonrimming steels in which the natural shrinkage occurring during ingot solidification is offset to an appreciable degree by gas formation. Accordingly, this type of steel includes the entire range intermediate between rimmed and fully killed steel. The resultant ingot structures range from those of steels that have been capped to suppress a weak rimming action to those steels that have been killed in the furnace or ladle to the extent that no mold deoxidizer is required. A survey reported to the 1946 AIME Open Hearth Conference' reviewed the practices being used at 28 major plants for the deoxidation of semikilled steel. It was found that most plants divide the deoxidizing additions between the ladle and the mold many of the plants adding only minor amounts of deoxidizer to the bath or none at all. Silicon, aluminum and titanium were the only elements used for ladle and mold deoxidation. Mold deoxidizers were added as capping additions or as uniformly fed additions. Several practices were reported in which the steel was deoxidized in the ladle to the extent that little or no mold deoxidation was needed. No reference was made to grades capped by using special mold designs or special closures over the top of molds. Accordingly, specific examples of this type of practice will not be considered in this paper. It appeared from this survey of deoxidation practice that semikilled steel can be divided into three types, namely, capped semikilled, in which aluminum is added near the end of pour or after shut-off, intermediate semikilled steels, in which aluminum is fed uniformly throughout the pour, and semikilled steels requiring no mold deoxidation. In an attempt to cover the ingot structures obtained by the range of practices reported in the preceding survey, three series of experimental heats were made using increasing amounts of aluminum, silicon and titanium for ladle deoxidizers. The ladle deoxidizers were added as stick aluminum, 50 pct ferrosilicon and 20 pct ferrotitanium (4 pct carbon), respectively. Two methods of studying ingot structures were used. Ingots were removed from production, burned longitudinally and machined to allow observation of the central pipe and coarse blowholes. During the machining operation much of the finely porous subsurface structure was obliterated. Accordingly, observations were also made

Jan 1, 1947

By C. L. Mantell

When considered from the viewpoint of world annual output, tin is one of the rarest metals. Its annual production is exceeded by that of iron, copper, lead, zinc. aluminum, magnesium; probably by that of the ferroalloys of manganese, chromium, and silicon, and possibly by the world production of sodium; while it exceeds only the output of such common metals as nickel, antimony, mercury, tungsten, molybdenum, silver, gold and the precious metals. Cassiterite, or tinstone, until the second World War was the only mineral that was an important source of tin. This mineral is frequently called "tin ore." The use of the term should be restricted to the ore containing the mineral, and not, in addition, to the concentrates of the mineral obtained from the ore or from stanniferous alluvial deposits. There is a general misuse of the term "tin ore" in the Malay Peninsula and adjacent localities, particularly where most of the mineral is obtained from secondary stanniferous deposits. The term "cassit-erite" is exclusively used in mineralogical, geological, and other scientific writings; it is unfortunate that it is not mole frequently used when referring to the occurrence of the mineral on an economic scale. Ores "Tinstone" js a very convenient old English term deserving a more frequent use. Tinstone is a dioxide of tin, or stannic oxide. When chemically Pure, as in the very rare transparent variety, it has a metallic content of 78.6 per cent tin. Frequently, however, the crystals and grains contain appreciable amounts of impurities, chiefly iron and tantalum. The impurities here referred to are those actually in the mineral itself. Tinstone usually has a deep brown or black color with an adamantine luster. Several other colored varieties are known, among them red ruby tin, yellow rosin tin, and yellow wax tin. the names jn each case being descriptive of the mineral's appearance. Ainalite is a variety of cassiterite con-taining almost g per cent of tantalum pentoxide. Sparable tin, tooth tin, and needle tin, as a result of their acute dite-tragonal pyramidal crystalline form, re-ceive their names from their crystallo-graphic appearance. Wood tin is a compact variety of cassiterite composed of radiating fibers resembling dry wood. Toad's-eye tin is a similar variety on a smaller scale in which the fibers appear to resemble the eye of a toad. Stream tin is water-worn tin-stone. Float tin is sometimes employed to describe the cassiterite occurring in soil derived from the weathered surface of a mineralized area. Stannite is an ore of tin of lesser impor-tance than cassiterite. It is a sulphide of tin, copper, and iron, sometimes known as tin pyrites or bell-metal ore. Its chemical com-position is sometimes expressed by the for-mula Cu,S.FeS.SnSz with zinc usually present in varying quantities. The tin con-tent varies from 22 to 27 per cent and there is about 29 per cent of copper, 13 per cent of iron, and 30 Per cent of sulphur.

Jan 1, 1944

By R. W. Raymond

Four portraits have recently been hung in the rooms of the Institute, in recognition of four illustrious men with whom we, as mining engineers and metallurgists, may claim fellowship. Luther. Martin Luther is often spoken of as a miner's son; but it is less generally known that John Luther, his father, was, in a small way, a metallurgist as well as a miner, and that Martin Luther himself, in his youth, worked in a forge or bloomary established by his father at Mansfeld, Saxony. Since this is the only manual occupation on his part of which we have ally record, it is fair to assume that the great scholar and reformer derived from his work in this department the practical familiarity with the conditions of human labor so abundantly evinced in the homely illustrations which accompany even his serious arguments, and in the witty apothegms of his famous " table-talk." At the same time, I must confess that I have not been able to find in any of the works of Luther distinct allusions to mining or metallurgy. Körner. Theodor Körner, a student at the Mining Academy of Freiberg, Saxony, threw himself into the German War of Freedom in the first decade of the 19th Century. He was a poet of no mean rank; and his " Song of the Sword " is one of the great lyrics of literature. This poem was written but a few days before he died in battle. In many respects, Körner may be compared with our American hero Theodore Winthrop, who similarly met a glorious fate, and left behind him prophetic indications of geniue, of which his premature death prevented the complete fulfillment. I am informed by a member of the Institute that Körner's admirable collection of minerals is still preserved at Dresden.

Jan 1, 1909

By Oswald J. Heinrich

The deposit of rock salt along the shores of Lake Huron, in Canada, has been brought before the public during the last six months, in consequence of the developments made by the diamond drill at Goderich. It is likely to be of more than common interest in the future, since now actual operations have been commenced to exploit the deposit and work it by regular mining. The methods of mining rock salt being but little known in this country, it may be of interest to make a few general remarks in regard to them before entering, upon the description of this new mining enterprise. According to the conditions under which salt is .found in nature, various methods have been resorted to, in course of time, to make it available for the human race. They consist in: A. Natural or solar evaporation. B. Artificial evaporation, either exclusively or in connection with the former. C. Mining the rock salt in its natural state, and converting it to a marketable product, if need be, by simply pulverizing it. Without going into the details of the various methods of making salt available in an economical point of view, dependent on its degree of pority, or on various local influences, we cannot refrain from alluding to the general principles which influence the mining and manufacturing of salt in countries where they have been carried on for hundreds of years. ECONOMICAL CONDITIONS FOR SUCCESS. Salt being one of the articles of prime necessity, and one universally distributed throughout nature, it must always be sold at a comparatively low price. It would even be quoted in the market at prices far below those now ruling, if it were not for the high duties generally imposed upon it, for purposes of state revenues. But looking closer into the subject, it will be found that this duty is probably one of the least oppressive, although in some countries it exceeds sixfold the cost of profitable production. To produce the article at the lowest possible cost requires, therefore, a

Jan 1, 1879

[A discussion suggested by the paper of Mr. James Gsyley on " The Development of American Blast-Furnaces," read at the New York meeting of the Iron and Steel Institute, October, 1890, and reprinted from the Journal of that Society in Bans. A. I. M. E., xix., 932.1 (Cleveland Meeting, June, 1891.) [Kote by the SecretaRy.—The three following contributions to this discussion were communicated by letter to the Secretary of the Iron and Steel Institute, and were received by me too late for insertion in vol. xix. of the Transactions.] FRed. W. Gordon, Philadelphia, Pa.: Mr. Gayley's paper calls our attention to the very expensive blunder hitherto made in blowing too much wind in a given size of furnace. It was quite evident from an analysis of blast-furnace operations that the work we are doing in America is anything but creditable. We were consuming fully 1000 pounds of coke more than is necessary for each ton of iron made, and were hammering the linings of our furnaces out in a very short time, and nothing but the very high price of pig-iron would stand the drain. It had long been my pet scheme to gradually reduce the quantity of air, at the same time increasing the burden to the maximum and decreasing the limestone, and determining, if possible, the fuel economy which might be reached consistently with the maximum production of the furnace. After the success of the Isabella furnace at Pittsburgh, which was remodelled in 1880, the North Chicago Rolling Mill Company had

Jan 1, 1892

By H. D. Outmans

The paper presents a theoretical approach to the drilling problem based on rock mechanics and drilling fluid hydraulics at the bottom of the hole. The volume of the fractured rock around the vevtically penetrating bit tooth is expressed as a function of the tooth pressure, included tooth angle, drilling fluid pressure and formation characteristics on the assumption that fracturing is preceded by plastic flow. This assumption is verified by comparing the stress-strain state in the rock around the tooth to that in equivalent tri-axial compression tests described in the literature. An expression similar to the above is derived for the tooth penetration in the cuttings which have been retrained at the bottom. Combination of these two expressions leads to a relation between depth of tooth penetration in the virgin rock, the amount of cuttings retained and the other variables. By relating the cutting volume to the drilling rate, the hydraulic bit horsepower and the drilling fluid pressure, it is possible to arrive at formulas which express the drilling rate for viscous or turbulent flow at the cutting surface, for instant clearance, for retained cuttings, or for false tooth foundering as functions of such variables as the drilling fluid pressure, hydraulic horsepower at the bit, rotary speed and bit weight. Plotting the results as drilling curves and comparing them to published field and laboratory data justify the conclusion that the drilling model presented in this paper approximates the actual mechanism. The drilling curves serve to explain some characteristic features of the rotary drilling operation, and the equations may be used for numerical evaluation of the effect of changes in the magnitude of some of the variables on the drilling rate. INTRODUCTION Published laboratory data have established that consistent and relatively simple relations exist between the drilling rate and any one of the important variables. For instance, the drilling rate usually increases at an increasing rate with the bit weight. It increases at a decreasing rate with the rotary speed, decreases at a decreasing rate with increased drilling fluid pressure and plots as an S-shaped curve against the hydraulic horsepower at the bit. This relative simplicity seems difficult to understand considering that the drilling mechanism is a continuous process of generation and removal of innumerable individual cuttings, under different conditions for each. On the other hand, the drilling rate (although determined by this complicated process) only expresses the average rate at which cuttings are produced and cleared away. Ignorance of the drilling mechanism as far as individual cuttings are concerned, therefore, does not exclude the possibility of arriving at a concept of the drilling rate provided the mechanism is properly evaluated for the average cutting. It then may be concluded from the first sentence that this concept also should be rather simple in its final form, although not necessarily in its derivation. Two aspects of the drilling mechanism are difficult to evaluate — (1) the effect a penetrating drilling fluid has on the stresses around a bit tooth entering the formation and (2) the volume of rock fractured by a penetrating tooth which moves both vertically and horizontally. Thus, the scope of this paper has been limited to an evaluation of the drilling mechanism in formations which have a low permeability and are drilled with hard-formation bits. In the following sections, the drilling mechanism is analyzed in three steps. First, the effect of tooth pressure on the volume of fractured rock around the tooth is evaluated, assuming initially that the bottom of the hole is free of cuttings and, then, determining the effect of a layer of these cuttings. In the second section, the thickness of this layer is related to the drilling rate and the chip clearance time, and the latter to the hydraulic horsepower at the bit under conditions of laminar and viscous flow at the cutting surface. In the third section. the previous results are combined with the rotary speed into drilling equations which are then plotted and compared to experimental drilling curves. This section also contains some numerical examples of drilling rate computations. Some mathematical details are given in the appendixes. Finally, it should be mentioned that, although some drilling equations are already available, these equations (with the exception of one' which attempts to relate drilling rate to drilling fluid pressure) are all mathematical expressions for observed laboratory data; none considers the effect of delayed drill-cutting removal, an essential feature of the drilling mechanism.

By L. G. Sprague

The fact that this latest and most modern of the Universal Atlas Cement Company's plants at Northampton, Pa., is the fifth to be built on these same properties, and their development has been coincidental with the development of the portland cement industry in the United States, makes it appropriate to sketch briefly the history of the successive steps. In 1889 the original Atlas Portland Cement Co. started its first plant at Coplay, Pa., which is just across the Lehigh River from the present operation. Here the first successful rotary kiln was operated. These kilns were 6 ft. in diameter and 60 ft. long; they used oil for fuel until 1895, when the company developed the first burner for pulverized coal. Shortly thereafter the Atlas company acquired property on the Northampton side of the river and built its plants Nos. 2, 3 and 4 within the following 10 years. During the same period the production of portland cement in the country was increasing rapidly; owing largely to the development of the rotary kiln and improved grinding machinery, which lowered the cost and improved the quality of cement. The general tendency has been to increase the length of the kiln and grind both raw material and finished product Liner. In 1930 the Atlas Portland Cement Co. and the Universal Portland Cement Co. united to form the Universal Atlas Cement Co., and the same policy of improving the product has been continued. Because of the large amount of research carried on by manufacturers and consumers, including government agencies, there is a demand for cement of different chemical and physical characteristics for different kinds of structures. Hence many engineers in recent years have been asking for cements of certain definite chemical analyses, which they think are best for their particular jobs. This has compelled the manufacturer to make several different types of cement by variations of chemical composition. These are generally made by adding silica, iron and lime to the standard mixture, which necessarily increases the cost of the product. Therefore the Universal Atlas Cement Co., in planning the modernization of its Northampton plant, gave this particular phase of the problem considerable study. Research and experimentation have developed the practicability of making a wide variety of differently proportioned raw materials for cement by a process of subtraction rather than by addition. With such process, it is possible to discard the undesirable aluminates and silicates from the ample supply of impure limestone available from the company's property rather than by adding high-grade limestone obtained from distant sources. This procedure requires mineral separation by centrifugal sedimentation and froth flotation. Since the chemical and mineral constituents of Lehigh Valley "cement rock," together with the methods of separating them, have been thoroughly described in published

Jan 1, 1948

By Gregoire Gutzeit

THE development reported in this paper was begun by the author a number of years ago, while he was a lecturer on complex chemistry and metallurgy at the University of Geneva, Switzerland, and in charge of that institution's research laboratory. Completion of the development and reduction to practice did not occur until last year at the Westport mill, laboratories and testing plant of the Dorr Company, where this method is now in general use for identifying and localizing the mineral constituents of both metallic and nonmetallic ores. It has proved to be a practical tool of the metallurgical staff at the Westport mill, devising unusual milling flowsheets for complex and refractory ore bodies. In a number of instances the findings have been successfully used as fundamental factors in the design and specification of complete metallurgical plants. Seven examples of the use of this method in solving practical milling problems are included in this paper. GENERAL OUTLINE The microscopic study of an ore, at first only of scientific interest, has proved extremely useful both in practical geology and in ore dressing. The binocular magnifying lenses for rough observation of the products of an ore-dressing plant have been in use for a long time; but only within the last few years has a systematic microscopical examination of ore samples become a normal tool of any research laboratory dealing with ore-dressing problems. Such an investigation avoids considerable loss of time as well as empirical trials and useless experimentation. It will generally indicate the correct grinding and often will suggest the proper method of treatment. The petrographic determination of transparent minerals in thin sections, or even in grains, has been perfected, and very accurate methods have been devised to meet any requirement, such as Feodorof's rotating stage on the petrographic microscope. Almost all transparent minerals can be positively identified by their optical properties; i.e., isotropism, anisotropism, color, pleochroism, refractory index, cleavage, crystal form, birefringence, and others. The identification of opaque minerals is a much more difficult problem and can be done only by using polished specimens. Unfortunately, most of the valuable metallic ores, and even some of the nonmetallic ones, are not transparent; therefore it has been necessary to develop a special technique for the determination of these minerals. Some optical properties-for instance, isotropism, anisotropism, color, cleavage, reflecting power-are often adequate in roughly classifying the mineral. Only "reflecting power" is a measurable value, provided special photometric equipment (ocular photometer or Orcel's photoelectric cell) is available. The other optical properties mentioned are more or less relative,

Jan 1, 1942

By McQuiston. F. W., E. C. Tveter

Sulfide mineral flotation has been applied to all naturally occurring sulfides and several synthetics. Of even more importance is the fact that economic separations are now being made between sulfides of essentially all the different metals. The variety of methods and minerals involved in such practice make it difficult to characterize them except in terms of general factors which influence the practices used. Ore grade and mineral composition are the major factors determining the processing method and the extent to which processing must be carried to achieve optimum economic return for a given ore. Since the choice of treatment is also affected by such factors as mine location, shipping charges, smelter schedules, each ore is a separate problem even though mineralization may be similar; The dissemination of minerals in the ore naturally determines the fineness of grind required for liberation. However, since grinding costs are a major part of most milling costs, emphasis has been placed on means of rejecting gangue minerals at the coarsest size possible. This has led to increased flotation of middling products or bulk concentrates for subsequent regrind and retreatment. Reagent development has also played an important part in present plant practice. Relatively few new frothers and collectors have found application during the past 25 years and many have disappeared. In general practice, re- agent combination choice is now fairly consistent with the type of collector being chosen on the basis of the mineral to be floated, the frother on the basis of pulp alkalinity, and the modifiers primarily on the basis of minerals to be rejected. REPRESENTATIVE FLOTATION PLANTS The milling operations described in this chapter have been selected primarily to illustrate these differences in ores and processing methods for the major types of ore minerals. The compositions of these minerals are shown in Table 1. These concentrators were chosen as representatives, both on the basis of geographic location and mineral content of the ore treated. At least one mill is included from each of the major producers, but the primary object was to show the practice employed for the various ore minerals and the different approaches to the middling problems encountered. The individual plants which have contributed data to this study are listed in Table 2, with a summary of their operating data, and are discussed in the order shown.

Jan 1, 1962

By L. G. Sprague

The fact that this latest and most modern of the Universal Atlas Cement Company's plants at Northampton, Pa., is the fifth to be built on these same properties, and their development has been coincidental with the development of the portland cement industry in the United States, makes it appropriate to sketch briefly the history of the successive steps. In 1889 the original Atlas Portland Cement Co. started its first plant at Coplay, Pa., which is just across the Lehigh River from the present operation. Here the first successful rotary kiln was operated. These kilns were 6 ft. in diameter and 60 ft. long; they used oil for fuel until 1895, when the company developed the first burner for pulverized coal. Shortly thereafter the Atlas company acquired property on the Northampton side of the river and built its plants Nos. 2, 3 and 4 within the following 10 years. During the same period the production of portland cement in the country was increasing rapidly; owing largely to the development of the rotary kiln and improved grinding machinery, which lowered the cost and improved the quality of cement. The general tendency has been to increase the length of the kiln and grind both raw material and finished product Liner. In 1930 the Atlas Portland Cement Co. and the Universal Portland Cement Co. united to form the Universal Atlas Cement Co., and the same policy of improving the product has been continued. Because of the large amount of research carried on by manufacturers and consumers, including government agencies, there is a demand for cement of different chemical and physical characteristics for different kinds of structures. Hence many engineers in recent years have been asking for cements of certain definite chemical analyses, which they think are best for their particular jobs. This has compelled the manufacturer to make several different types of cement by variations of chemical composition. These are generally made by adding silica, iron and lime to the standard mixture, which necessarily increases the cost of the product. Therefore the Universal Atlas Cement Co., in planning the modernization of its Northampton plant, gave this particular phase of the problem considerable study. Research and experimentation have developed the practicability of making a wide variety of differently proportioned raw materials for cement by a process of subtraction rather than by addition. With such process, it is possible to discard the undesirable aluminates and silicates from the ample supply of impure limestone available from the company's property rather than by adding high-grade limestone obtained from distant sources. This procedure requires mineral separation by centrifugal sedimentation and froth flotation. Since the chemical and mineral constituents of Lehigh Valley "cement rock," together with the methods of separating them, have been thoroughly described in published

Jan 1, 1948

By S. W. Parr, F. B. Hobart

Studies relating to the behavior of coal toward oxygen may have for their purpose the determination of the fundamental factors that underlie spontaneous combustion, weathering and deterioration, and the general topic of storage, as well as those properties that have to do with the underlying principles of carbonization. There is also involved the possibility of developing information relating to the constitution of coal, or at least the chemical properties inherent in the type substances of which the coal is composed. It is the purpose here to record certain results bearing on this latter idea of the chemical properties that characterize the main constituents of bituminous coal. The avidity of freshly mined coal for oxygen has long been known. What happens to the coal or to the oxygen is not so clear. Is the oxygen absorbed or chemically combined, and what are the conditions as to time, temperature, texture, or components that are most involved in the phenomenon? If we subject a weighed quantity of coal maintained at a constant temperature to a current of oxygen, we may obtain three factors of interest—the weight increase of the coal, the weight of the water, and the weight of the carbon dioxide formed in the reactions involved at that

Jan 1, 1925

By C. E. Birchenall, J. R. Stubbles

ThE equilibrium between lead, lead sulfide (PbS), and atmospheres of H2 and H2S has been determined previously by Jellinek and Deubel,' and SudO.' Blbem and Kroger3,4 used Jellinek's data to calculate the sulfur dissociation pressures of PbS crystals (of

Jan 1, 1960

By Bernard Kopelman

THE particle size of tungsten metal powder used to make tungsten wire for use in radio tubes and incandescent lamps must be closely controlled if the highly desirable feature of nonsagging is to be achieved. Our knowledge of factors affecting the particle size of tungsten powder has been limited by lack of any extensive literature on the subject. There are no publications in the English language discussing this problem in any but the most general terms. In this paper the published experimental information on this subject is reviewed. Based on the published data and some experimental data obtained by the author, a general theory of factors affecting the particle size of tungsten metal powder obtained by reduction of the oxides in hydrogen is offered. The various forms of the tungsten oxides and tungsten metal used in the experiments described were those prepared by com¬mercial processes, so that the experimental results obtained on a laboratory scale would be more directly applicable to actual commercial reduction conditions. The effects of time of reduction, temperature, and concentration of water vapor in the hydrogen were observed by counting microscopically the particle distribution obtained. Such a method is very time-consuming, but the data obtained by this method are more reliable and more fundamental than those obtained by any of the several instruments available for measurement of particle size. REVIEW OF THE LITERATURE Until the year 1912, tungsten powder was prepared commercially by reducing tungstic oxide in the presence of carbon or one of its compounds (such as glycerin or ethylene glycol), usually by firing the mixture in a closed fire-clay crucible. The resultant powder contained about 95 per cent tungsten.1 When reduction of tungsten oxides by H2 was introduced commercially, 2.1 several investigators became concerned with the equilibria between tungsten metal and its several oxides in an atmosphere of hydrogen and water vapor.4-13 While these many investigations were in fair agreement on the equilibria there arose considerable controversy as to the chemical composition of the oxide between W03 and W02, because of very poor agreement on the equilibrium constant of this intermediate oxide reduction reaction. The controversy on this intermediate oxide has been in progress for more than 10o years. Thus Malguti,14 in 1835, decided it was W02.56. In 1857, Riche" proposed the formula W308. Usler,16 in 1865, confirmed the existence of W308 and further proposed W4O11. In 1897, Desi17 agreed upon W308 but also found W5014.

Jan 1, 1946

By T. F. Bower, M. C. Flemings

Results are reported of a study of surface dendrilic structure of an Al- Cu alloy solidified against a chill wall. Most primary and secondary "arms " in the surface dendritic structure are arranged orthogonally, giving the impression of strong preferred orientalion on the surface. However, no such preferred orientation exists and it is therefore evident the arms do not represent (100) directions. The primary arms are shown to be interseclions of a (100) plane wilh the chill plane, or, equally often. the projeclion of a (100) direction on the chill plane. Secondary dendrite arms are usually within a few degrees of 90 deg to the primary arm, independent of grain orientalion. Prirary, secondary, and higher-order surface dendrite arms almost always represenl intersections of (100) platzes with the chill surace, or pvojections of (100) direclions. Growlh of secondary arms is favored on the side of the primary arm where a (100) direclion points toward the chill surfAce a1 a Lou, angle. Surface dendrile arms are often observed to be bent. In these cases, the crystal lallice changes orientation; bending is concave to the chill surface. In a previous paper,' a technique was discussed whereby large grains can be obtained at a chill surface. The technique used involves quickly drawing superheated liquid A1-4.5 pct Cu alloy into a thin copper mold, so that the mold is full well before solidification begins. The chill surfaces employed are polished copper blocks coated with amorphous carbon. Shrinkage during solidification between dendrite arms and grains delineates both, without the need for polishing or etching of the cast surface. The grain structure of the chill surface was discussed in a previous paper;' in this paper, the dendrite arms within each grain are examined. Previous work on surface dendrites includes that of Edmunds, who studied the development of preferred orientation in zinc, cadmium, and magnesium.' In zinc and cadmium, he found that the surface region has a (0001) texture (parallel to the chill surface). Walton and Chalmers reasoned that, since the fast growth (1010) directions are in the basal plane, nuclei which have this plane parallel to the mold wall would produce larger grains than nuclei with other orientations. Hence, the texture observed is as expected.3 The same authors, in measurements on aluminum ingots, found no preferred orientation at the mold wall. However, the X-ray technique they used measured the preferred orientation in terms of grain numbers, not grain areas; larger grains were weighted equally with small ones. No preferred orientation is expected on this basis at the chill surface. In a later paper,' Edmunds stated that experiments show a random grain orientation at the surface in die cast aluminum; his technique, also used in his earlier paper, takes account of grain area. Little work has been published on the dendritic structure of metal chill grains. Recent work of Biloni and Chalmers on "predendritic growth" shows the change in morphology from spherical to dendritic during the initial stages of freezing, 5 but this work did not include detailed examination of the fully developed dendrites. Other pertinent work includes that of Lin-denmeyer, who investigated the growth of ice dendrites. 6 When growth was on a substrate, the dendrite axes were bent. The bend corresponded to a change in orientation of the crystal lattice and occurred in such a way as to align the basal plane to the substrate. DENDRITE STRUCTURE Fig. 1 shows the chill surface of a typical casting poured above the critical temperature necessary to produce coarse grains. A cursory examination of these grains shows that the surface dendrite arms within most of the grains are oriented roughly perpendicular to each other. One is tempted to assume that these are (100) directions and that, therefore, marked preferred orientation exists at the chill face. This, however, is not the case. Each of the grains in the casting of Fig. 1 was separately identified, Fig. 2, and its orientation determined by the Laue back-reflect ion method. Results are given in Fig. 3 and it is seen there that no preferred orientation exists. Even when grain area is accounted for, there is no significant preferred orientation. The relationship between surface grain structure and crystal orientation was then obtained by assigning X and Y axes to the casting surface, Fig. 1, and assigning the same axes to the stereographic projections of each grain. Thus, the visible surface structure could be compared readily with grain orientation. This was done for fifty-five of the grains of Fig. 1. Results of this study on three typical grains are described below, and some general observations given subsequently. Fig. 4 shows the structure and stereographic projection of a grain which lies near the (100) zone (with respect to the casting surface). The X and Y directions are marked on the projection, and the photomicrograph mounted with the same orientation. Poles of the stereographic projection represent crys-tallographic directions in the grain which point out of the casting, toward the chill. Two (100) directions are shown in Fig. 4. A line joining the center of the projection and a pole represents the projection of the pole onto the X-Y plane (chill surface). Two such lines are shown in Fig. 4 (solid lines). A line joining the intersection of a great circle with the circumference of the projection gives the trace of a crystallo-graphic plane in the chill surface; two such traces are shown (dashed lines).

Jan 1, 1968

By E. B. Olds, Burt. B. P.

THE mines of the Consolidated Coppermines Corporation are at Kimberly, in the Robinson mining district, White Pine County, Nevada. [ ] The blanket-like ore body is a disseminated copper deposit occurring in monzonite porphyry and, in general, is thoroughly fractured, altered, and soft. Contained in the porphyry ores of the extreme westerly end of the ore body are pendants of hard garnetized limestone. A younger monzonite porphyry underlies the ore-bearing porphyry, which in turn rests on limestone and shale. Block caving is the standard method of mining. Because of the undulating top and bottom contours of the ore body, the main haulage level was selected and driven at an elevation calculated to be the most economic depth below the major tonnage sections of ore. Fig. I shows a general section of the Emma Nevada ore body taken through its long axis, which lies roughly in an east-west direction. Mine openings in either type of porphyry require heavy timbering and constant maintenance and repair, particularly in horizons at or near the ore body contact. Where the ore lies close to the haulage level, slusher hoists and scrapers are used to transport horizontally the ore to the main haulage drift (Fig. 2). For the extraction of high-lying ores, a system of branch raises is used; and the ore is mined by gravity. Slushing methods are used also on the high-lying ore-body fringes. Slusher drifts are often driven from the fringe branch raises that have already served their purpose, in order to minimize the cost of the preparatory work in mining the thinner portions of the ore body (Figs. 3 and 4). When a block of ore is developed and undercut, it is economical to mine it as rapidly as the ore will cave efficiently; and this procedure often presents safety problems because of men with equipment working in a relatively congested area. The single-track main haulage drift is about 1 ½ miles long, and lateral haulage drifts are driven under the ore body at intervals determined by the height of the ore body above the haulage level. In order to minimize repair and maintenance, the

Jan 1, 1946

By George H. Eldridge

The existence of phosphate of lime within the State of Florida has been known for over a decade; but until the spring of 1887, the extent and value of its deposit.;, possibly with one exception, were unsuspected. In that year Col. T. S. Moorehead, induced by a statement made by Mr. J. F. Le Baron, examined the gravels of the Peace river, recognized the importance of the deposits, acquired such tracts as were regarded as desirable, and established the pebble-phosphate industry, making his first shipment in May, 1888, to the Scott Manufacturing Company, of Atlanta, Ga. The establishment of the pebble-phosphate industry upon Peace river was followed, a year later, by the discovery in Marion county of an entirely different class of phosphate, now commonly known as hard-rock phosphate. This discovery was made early in June, 1889, by Mr. Albertus Voght, in a well which he was sinking upon his place near the site of the present town of Donnellon, about 20 miles southwest of Ocala. To it, more than to any other cause, is due the rapid and remarkable development of the present great industry of the State. It was, however, Mr. John F. Dunn, of Ocala, who turned to advantage the accidental discovery of Mr. Voght. After acquiring a knowledge as to the chemical composition and value of the material, Mr. Dunn and hid associates purchased several thousand acres of land and organized the Dunnellon Phosphate Company for active operations. From Dunnellon as a center, prospecting rapidly extended in all directions, until at the present day important mines have been opened over a well-defined tract more than 200 miles long and about 6 miles wide. This tract lies par-

Jan 1, 1893

By R. B. Brinsmade

The following is an attempt to form a formula by which a mine can be quickly evaluated, aft,er all pertinent physical data have been collected from observations on the ground by a competent mining engineer. Assumptions. Let Q = price to be paid for the mining property. M = cost of developing the mining property to yield y tons of ore daily. P = cost of suitable plant to treat y tons of ore daily, p = value of said plant when the mine has been exhausted. C = total fixed capital investment. W = working capital investment. Then let C + W = total capital investment. y = required yield of ore daily in tons. Y = yield of ore yearly in tons, d = number of producing days per year, u = average operating profit per ton of ore. R = rate of interest to be earned on total investment of (C + W). r = rate of interest to be earned on sinking fund annuity. v = tons of positive ore available in mine. x = tons of probable ore available in mine, z = tons of possible ore available in mine, m = fractional factor to change probable to positive ore. n = fractional factor to change possible to positive ore. Q = tons of total ore available in mine. t = time in years to exhaust mine at rate Y.

Jan 1, 1909