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By J. Parke Channing

THESE days of great industrial and social problems in America produce many suggested solutions and great changes. The practical engineer and employer of labor views these problems differently from the labor leader or the social reformer, but as never before he is sincerely interested in solving them in a way that will be just to all. The inevitable tendency of the day is toward "industrial betterment," "safety," "industrial education," "efficiency," and the many other things' which have become so familiar to progressive employers. There is no longer any question that these things are worth while from both the hu¬man and economic standpoints. They "pay" in dollars and cents. The very center of final success in improving conditions and increasing the efficiency of workingmen must be the spirit of fairness and a knowledge on the part of the employer of how to deal sympathetically and intelligently with his employees. Every progressive employer knows how greatly he desires foremen, superintendents, managers and others who possess these qualities. On the other hand, we are all familiar with serious mistakes made by young graduates of engineering schools who have had no opportunity to develop these qualities, and who have no real appreciation of the worth of the workers. Indeed, one wonders whether much ill-feeling, labor difficulties, and many strikes could not be avoided if such men had the right attitude. Is there any way of remedying this condition? If this particular difficulty can be solved, if these young engineers, many of whom are our coming leaders of industry, can be given the right perspective and the right understanding of these other problems in addition to fair, sympathetic methods of handling men, many of our other problems will be solved-not at once, but gradually and permanently, as these men make good and become influential in paths of industrial righteousness and industrial peace. Many progressive employers of to-day have enlarged their own perspective and realize the great importance of enlarging the perspective of those who shall follow them.

Jan 3, 1915

Asarco's Sacaton unit, located near Casa Grande 45 miles southeast of Phoenix, Arizona, was brought into production in February 1974. Ore reserves are estimated at about 48 million tons at 0. 95% copper. Mineralization of the orebody comprises both sulfides and oxides, with minor amounts of gold and silver. Preproduction stripping of 40 million tons of overburden was accomplished in less than two years. Mining is done at the rate of 9,000 stpd ore with a waste-to-ore ratio of 4.6 to 1. 0. The mine is developed on 40 ft benches, and by 1979 underground block caving methods will be phased into the operation at a depth of 1900 ft. Ore is mined in the pit two shifts per day, seven days per week, and is hauled with 85 ton trucks to the primary crushing plant. Trucks dump their loads into a 165 st capacity pocket equipped with a chain curtain which controls the ore flow to the 6 ft 6 in. wide by 24 ft long (2. 0 m by 7.3 m) vibrating feeder grizzly. Grizzly bars are inclined at a 5º slope on the first half of the 0 feeder and 10 on the last half. Grizzly bars have 6 in. (13 mm) spacings which pass 60 to 70% of the mine-run ore. Grizzly oversize drops into a 48 in. by 60 in. (1. 22 m by 1.52 m) primary jaw crusher. Both the grizzly undersize and crusher discharge are collected on a 42 in. (1.07 m) conveyor belt operating 398 fpm (121 mpm) for transport to the 4,750 live st intermediate storage area ahead of the fine crushing plant. Capacity of the Sacaton primary crushing plant is a nominal 1150 stph (1043 mtph). It operates fourteen shifts per week with an effective crushing time of four hours per shift. An economical open-air design was used, but an enclosed operator control room is located on the upper level where the dump truck, feeder, and crusher can be observed. Oversize rocks are removed from the feeder or crusher by the 10-ton bridge crane. Direct operating costs for this well-designed primary crushing plant, including labor, power, supplies, and maintenance, are only 3.5 cents per short ton. An interesting feature of the Sacaton orebody is that the non-sulfide copper minerals, which are less prevalent than the sulfides, occur intimately with the sulfide minerals; for this reason, all ores as mined are treated in a flotation circuit using sulfidizing methods for oxide recovery.

Jan 1, 1978

By John B. Mertie

More than 200 minerals are known that contain the rare-earths and thorium. Monazite and bastnaesite, however, are the principal commercial sources of the rare-earths, and monazite is the principal source of thorium. Monazite was in demand in earlier years mainly for its content of thorium, which was used in the manufacture of Welsbach gas mantles; but beginning in the latter part of the nineteenth century, the rare-earths became progressively more important than thorium. Recently, however, the tenor of thorium in monazite has again become a matter of commercial interest. Rare-Earths and Thorium The rare-earths proper are defined by spectroscopists as the oxides of 14 elements with atomic numbers ranging from 58 to 71; but more commonly lanthanum, with an atomic number of 57, is also included, in which case these 15 elements are often designated as the lanthanides. Chemists also include scandium and yttrium as elements of the rare-earths. A second group of elements, beginning with actinium and including thorium, protactinium, uranium, and all the transuranium elements so far produced, with atomic numbers ranging from 89 upward, are chemically and structurally related to the lanthanides, and are known as the higher rare-earths or actinides. Berzelius discovered that the lanthanides could be separated roughly into three groups, based upon the solubility of their double salts, R2(SO4)3-K2SO4-2H2O, in excess of K2SO4. This separation, with the substitution of sodium for potassium, is still used commercially. The double salts of the cerium earths were found to be almost insoluble; the terbium earths, partly soluble; and the yttrium earths, completely soluble. Based upon the degree of solubility, scandium belongs with the cerium earths; and yttrium belongs with, and gives its name to, the yttrium earths. These three groups with their atomic numbers are shown below. [ ] The prevailing state of oxidation for the rare-earths is R2O3, but cerium may readily be oxidized to the tetravalent state, in which condition it may be separated chemically from the other cerium earths. Praseodymium and terbium may also exist with the valence IV, and samarium, europium, and ytterbium may have the valence II. Most of the elements of the rare-earths, however, are very difficult to separate, and their initial isolation required the work of many chemists over a period of 150 years.20 Fractional crystallization, fractional precipitation, and ion exchange are the principal methods so far used, and fractional precipitation is now employed commercially on a large scale. Ion exchange columns are

Jan 1, 1960

By William Metcalf

IT is the object of this paper to oppose unnecessary changes, and the introduction of new and confusing terms. From the earliest times of which we have any record on the subject, iron has been divided into general classes. Whether iron was first known in the form of wrought iron or of steel, is a matter of no importance; but at the time when cast iron first became a prominent article of commerce, wrought iron and steel were old products¬steel, at that time, being generally made from wrought iron by the process of cementation. As steel was then properly defined as iron containing more carbon than wrought iron, so cast iron was appropriately defined as iron containing more carbon than steel. Until the year 1776, when Huntsman first began the manufacture of a new kind of iron by melting wrought iron and steel in crucibles, the above definitions were all that could be desired. At first, probably, owing to imperfect appliances, only a highly carburized iron was melted in crucibles, and this led to its being called cast steel. If our ancestors had found a name for this material not involving the use of the word steel, we would have been saved the trouble of the present discussion. As the arts developed, new uses were found for steel, and the products of the crucible were varied, until at the present day it is quite common to find steel containing no more carbon than average wrought iron, and other steel containing nearly or quite as much carbon as some cast iron. All of these various products of the crucible have retained, or received, the name cast steel, usually abbreviated to the simple word steel, without causing any confusion among producers or consumers in regard to terms or their meaning. In 1855, twenty-one years ago, the Bessemer process was introduced ; this was followed in a few years by the Siemens-Martin process, and the products of both were at once, and have ever since been, called steel, from the fact that they were identical in chemical and physical properties with crucible cast steel. At the present time, then, we have three subdivisions of cast steel, viz.: Crucible cast steel, Bessemer cast steel, and Siemens-Martin,

Jan 1, 1877

By Donald W. Gentry, John F. Abel

This paper presents some of the results of a rock mechanics instrumentation program designed for, and implemented at. longwall panel 4N at the York Canyon Mine near Raton, New Mexico. The objective of this particular instrumentation program was to determine the response of the rock mass to longwall mining of a thick coal seam utilizing shield-type supports. Measurements were made in the rock along the ground surface, above the coal seam, in the coal seam, and immediately below the coal seam. Longwall test panel 4N had a total developed length of 762 m (2500 ft) and a face width of 168 m (550 ft), [Fig. 1]. The seam mining height in the panel area was approximately 3 m (10 ft). The average depth of overburden above the panel is 107 m (350 ft) although it ranges from 73-134 m (240-440 ft). Roof support at the face was provided by 116 Hemscheidt 320 HSL shields. These shield-type supports weigh 11.4 metric tons each.

Jan 1, 1979

From the early part of the thirteenth century bituminous coal was called "sea coal" in England, from the fact that the coal seams in the Fife and Northumberland fields outcropped on the shores of the North Sea and wave action and erosion resulted in lumps of coal being found in large quantities along the shores. These were gathered by women and children and sold for fuel to replace wood. Coal first reached London by sea, and there was a "Secole Lane" in that city in 1228. The term was gradually applied to all coal and for some hundreds of years all coal in England was called "sea coal", and the term is not yet obsolete there. Another reason for the use of a defining word was the fact that from time immemorial, in England and elsewhere, the word coal meant char-coal, or coal made from wood, as this was the common fuel of those times when anything other than natural wood was burned. The biblical references to coal all mean wood-coals or charcoal, and one must be careful in writings prior to at least 1750, when "coal" is used to be sure that it means stone, or mineral coal, rather than wood or charcoal. "Sea-coal" was in common use along the Atlantic seaboard until after about 1780; the first map showing coal in Pennsylvania, Ohip and Kentucky, made as of 1750, used the term "sea-coal here."2 In the interior stone coal was used by the early writers almost entirely, partly to differentiate it from charcoal, but largely because it was found with stones and rocks, and in most places then had the same value. "Fossil" coal was used by Hamilton in 1790. Jefferson called it mineral coal and pit coal in 1781. The term mineral coal was usually used in the governmental references to it as late as 1850, but the present use of bituminous coal began as soon as anthracite came on the market in 1820. Both mineral and stone coal still survive in the laws of several states specifying the weight of a bushel of coal.

Jan 1, 1942

By John B. Hastings

Chief among the varied problems facing the mine-manager is that of vein-structure and origin, which is highly important as a guide to successful discovery and development. If metalliferous deposits can be traced to the intrusion of waters along definite lines, then is there something tangible for him to study. But when he is told by a geologist that metalliferous deposits are due to some other cause, like magmatic differentiation, while not perhaps discouraged, he is impressed with a vague sense of new worlds to be conquered. In 1897, while manager of the War Eagle mine at Rossland, B. C., where, previously, the ore-bodies had not been supposed to occur as veins, I suggested to visiting members of the Canadian Geological Survey that, perhaps, the structure of the pyrrhotite-bodies of that mine, which had been proved to occur along well-defined fractures, might throw some light on the similar deposits of Sudbury, then held to be basic aggregations from the original magma. Later investigation has at least provoked discussion of this last deduction, questioning its correctness and suggesting that, perhaps, the deposits are actual veins. The following extract from a paper by J. E. Spurr' is quoted, because it sets forth Mr. Spurr's theory of the quartz-occurrences of the Silver Peak mines, which will be discussed in this article. It also graphically describes the attendant geological features. "The Drinkwater group of mines, which is the most important part of the Blair gold properties, and which has produced practically all of the million dollars' worth of ore, as above stated, may be taken as typical of the gold-veins which, though widespread and numerons, show a wonderful similarity of character. On the surface two adjacent veins outcrop, the Crowning Glory and the

Jan 1, 1906

By R. D. Boddorff, R. L. Ash, C. T. Butler, W. W. Kay

DRILLING and blasting in anthracite open-pit mines is a continuous problem to contractors and explosive engineers because of the diverse conditions caused by the nature of the geological formations, the extensive mining of the portions of coal beds near the surface, and the proximity of many strip pits to populated areas. Pennsylvania anthracite occurs in four separate long and narrow fields totaling only 480 sq miles. The coal measures are rock strata and coal beds that are considerably folded and faulted. The crests of the anticlines are eroded extensively. The beds outcrop on the mountain sides and dip under the valleys. At first only the upper portions of the synclines could be stripped. Now stripping to increasingly greater depths is economically possible, as is indicated by the fact that the proportion of freshly mined anthracite produced by strip mining has increased from 3.7 pct of the total tonnage in 1930 to 29.6 pct in 1950. Much of the rock overlying the deeper beds now being stripped is so extensively broken that considerable difficulty is experienced in drilling satisfactory blast holes and in using explosives in such manner as to insure a uniformly broken material easily removed by the excavating machinery. Such breaking of rock strata has occurred because the bed now being stripped has been mined extensively in former years by underground methods, and tops of gangways and chambers have subsequently failed. Draglines are used to uncover coal where the overburden can be moved with little or no rehandling. These machines range in size from those having a 2 cu yd capacity bucket on a 60-ft boom to those handling a 25 cu yd bucket on a 200-ft boom. Draglines are also used to strip to the bottom of the coal basins if the depth and the distance between the crops are not too great. For this type of operation blast holes are drilled full depth to the bed. These holes are commonly 30 to 90 ft deep; however, in exceptional cases, holes may be as shallow as 12 ft or as deep as 130 ft. Drilling is normally done for blasts of 12,000 to 60,000 cu yd of overburden, 30,000 cu yd being considered an average blast if vibration is not the controlling factor. Where the stripping of wide basins or the exposure of a moderately pitching vein makes the use of draglines impractical, dipper front shovels equipped with 4 to 6 1/2 cu yd buckets load into trucks. Overburden is removed in benches of 25 to 30 ft with blast holes drilled 4 or 5 ft deeper than the planned floor of the bench. For shovels under 5 cu yd bucket capacity the volume blasted varies from 8000 to 12,000 cu yd, whereas a volume of 30,000 to 50,000 cu yd of overburden is frequently blasted at one time for the larger shovels where vibration is not an important factor. During the past decade the churn drill, generally the Model 42-T Bucyrus-Erie blast hole drill equipped for drilling 9-in. diam holes, has become the most common blast hole drilling machine. Electricity powers half the churn drills in use and is preferred on the large strippings where electric shovels are operated and the working area is concentrated. On these operations the cost of additional electricity for the drills is less than the cost of fuel to operate diesel units because of the existing large demand load of the excavating equipment. Moreover, electric motors start more easily in cold weather and generally are less expensive to maintain. Diesel driven units are employed where a higher degree of mobility. is required. The average drilling speed is 8 ft per hr, although in softer rocks a rate of 15 ft per hr is attained. Where rock is hard and strata is badly broken, drill speeds may ' be less than 2 ft per hr. Low drilling production results under these circumstances when loose material falling from the upper portion of the drill holes causes drill stems to be jammed. Rock formations vary so greatly in the region that a 9-in. diam churn drill bit may become dull after drilling only 2 ft or may drill satisfactorily for 56 ft; however, an average of 35 ft is usual in sandstone of medium hardness. Dull bits are hoisted to flat bed trucks by the sand line of the drill and are usually sharpened in the contractor's bit shop adjacent to the job. Care is generally taken to cover the thread end of the bit with a cap. To facilitate handling of bits around the drill, a heavy thread protector having an eye top is becoming more popular than the flat-top rubber or metal cap furnished with new bits. The 9-in. diam blast holes for a 25 to 30 ft bench are normally on 18x18 ft to 20x20 ft spacings, depending on the character of the overburden, although in broken ground 15x18 ft centers may be used to obtain better breakage and a more even bottom for the bench. The patterns of holes for shots

Jan 1, 1952

By Carl Samans

IN their recent complete review of the subject of age-hardening,1 R. F. Mehl and L. K. Jetter classify the main types of precipitation-hardening alloys under two headings, depending upon the nature of the precipitate and the character of the precipitation. In the alloys falling under the first heading, which display "continuous precipitation" and are typified by the aluminum-copper alloys, "precipitation seems to occur without much regard to grain boundaries and apparently at much the same rate at every point," although an acceleration in regions plas-tically deformed has definitely been established .2 In the second type of alloys the precipitation is "discontinuous," as in the copper-silver system for example, and "seems to start in restricted regions, primarily those contiguous with the grain boundaries, and to spread inward in an advanc-ing wave." The reasons for the occurrence of these two greatly different types of precipitation are unknown, as are those for many other reactions that take place in the solid state. In the present paper only the first or "continuous" type of precipita-tion will be dealt with, for which a complete crystallographic mechanism has been worked out for the process of precipitation from a supersaturated solid solution of copper in aluminum. Based upon this mechanism, an explanation of the various property changes found experimentally will be developed in the second paper of this series. In later papers "dis-continuous precipitation" in copper-silver alloys will be explained, and an effort will be made to rationalize the two processes and types of precipitation.

Jan 1, 1940

By J. J. Pelley

NOT being a scientist, an engineer or a metallurgist, I consider it a very great honor indeed to be asked to address the American Institute of Mining and Metallurgical Engineers. Your program indicates that any lack of intricate and technical knowledge on my part in the field of metals will be well taken care of by others. Therefore, I believe any remarks that I have to make might well be devoted to the practical side of metals in railroad service and to the general railroad situation as I view it today. The experiments now being conducted on the railroads with so-called streamlined, light-weight trains and cars would not have been possible had it not been for scientific and engineering research in the past that made these lighter weight metals available. The railroads now have approximately forty units of these new trains in operation. Experi-ence with them will no doubt be a guidepost for future construction in that direction. While the best known experiments are in the passenger service, they are being extended into the freight service. Within the past three years there was constructed for the Association of American Railroads a new standard box car of ordinary carbon steel, in which, by reason of superior design, weight was reduced by 3500 to 4500 lb. under that of existing cars of like capacity. The Association is now testing a box car that weighs approximately six or seven tons less than practically all of the box cars of like capacity in service today. If these tests are successful a very large field for the use of lighter weight metals will immediately be opened, provided the costs of such metals are not too high. The fact that various experiments are being carried on with other classes of equipment constructed with so-called light-weight and non-corrosive metals no doubt will have a considerable effect upon the rebuild-ing of old equipment. The question then will be: How much can we afford to spend on the rebuilding of an old car weighing six to seven tons more than a new car constructed with the latest available materials?

Jan 1, 1936

By J. A. Herrick

SOME time ago the writer needed a cheap, light, and portable hydraulic lift or crane, that would be universal in its application, and that might -be suspended from crane-arms, overhead tramways, beams, etc., and serve all the uses of ordinary hydraulic cranes. This was accomplished by the use of a closed cylinder of wrought-iron hydraulic pipe, provided at its top with the ordinary gland and stuffing-box, through the centre of which passed another similar pipe of slightly smaller diameter, forming in effect the ordinary plunger of the regular hydraulic crane. Over the top of this centre pipe or plunger or piston a disk of iron is placed with two arms projecting several inches from the rim, as may be required. Through these arms long iron rods pass, one on each. side, and are fastened in place by nuts screwed to a bearing immediately above and below the disk. These rods also pass through the enlarged head of the lift down through a casting which closes the bottom of the cylinder, which, at the same time, gives them a bearing, and they terminate at a point below the casting equal to the height of the lift of the crane. The rods at this point pass through a disk similar to the top disk and are similarly secured in place. A common hook arranged with a swivel passes through the centre of the disk, and on this hook the load is suspended. For a stationary hanging crane the apparatus is suspended by two rods passing through the head of the lift in similar manner to the two already described, at right angles to the first, and of sufficient length to let the first set play between them and rise to the full height which the plunger will allow. The second set of rods terminates in an iron disk similar to the one at the bottom of the first set, and a hook through this in turn suspends the entire apparatus. In order to adapt the lift to an overhead tramway, single or double, in place of the suspending rods, a projection or offset is arranged on each side of the head of the lift and a journal cast in, allowing the lift to be suspended on pivots arranged as follows

Jan 1, 1879

By M. E. Lombardi

THE cost of maintaining the production of an operating oil company is one of the most important, as well as one of the most difficult to estimate, of the various items which go to make up the total cost of producing oil. In the opinion of many operators the cost of drilling and developing new wells, whose production takes the place of the loss in production through decreased yield of the old wells, is considered an operating expense, for the reason that the total income-producing power of the property is not increased. The cost of these new wells may be added directly to production expense, or it may he charged to capital account, and this account depreciated to the same extent. In either case, after an oil company is on a satisfactory income-producing basis, the money for drilling new wells to maintain production must be taken from the income of the company. Three factors determine the cost of maintaining production: First, the rate of decline in yield of the wells; second, the cost of drilling and equipping new wells in terms of their production; third, the decrease in initial production of new wells as the property is drilled and gas pressure taken off the sands, etc. In order to check theoretical figures for the cost of maintaining production the writer, with the assistance of Reed Bush and Fred Tough, has compiled the following figures. It is necessary first of all to choose a period of time during which production for the field or territory studied remains fairly constant. The longer the period of time over which this condition prevails, the better. It is also necessary to eliminate or allow for any extraneous condition which may have affected the production returns, such as the shutting clown of a large number of wells, the bringing in of a well of exceptionally large yield, the opening up of virgin territory where the flush yield is greater than will he had from new wells in after years, etc.

Jan 9, 1915

By W. M. Mobley

THE coal industry in Alabama, centered in Birmingham, has pioneered coal-beneficiation practice in the United States. The nature of the coal seams and mining methods employed have necessitated use of equipment and methods best adapted to the local conditions. Although all of the 30 or more coal seams now being mined in the state receive beneficiation, only the three most important producers of coking coals, the Pratt, Mary Lee and Black Creek, are discussed in detail. Nature of the seams, washability characteristics and equipment used are given as examples of the extent of work and progress in this field. The coal seams in Alabama are relatively thin and pitching, making mining difficult and expensive. Most of them are characterized by thin partings of rock or shale in the seam itself, as well as laminations of bone and rash. Mechanical mining methods have increased the occlusions from roof and floor to such an extent that nearly all the coal mined in the state must receive some treatment before it can be put on the market. During the year ending Sept. 30, 1943,1 18,129,792 tons of coal were mined in the state and 14,364,177 tons, or 79.2 per cent, received treatment. Excluding from the total the 1,223,014 tons produced by small wagon mines and the hand-picked lump coal, the amount receiving beneficiation would be more than 90.00 per cent. This is, by far, greater than in any other coal-producing state. The earliest record of coal beneficiation in Alabama was the installation in 1890 of an experimental Lung washer by Alexander Cunningham, for the Sloss Iron and Steel Co. at its blast furnace in Birmingham .2 In 1894 the Jeffrey Manufacturing Co. introduced the English Robinson washer, a combination of a hydraulic classifier and centrifugal machine. Erskine Ramsay, who has contributed so much to the coal industry, saw its possibilities as well as its mechanical defects and soon had his inventive mind at work. He added a sludge tank of his own design, and many Robinson-Ramsay washers were installed in Alabama and other states. The Stewart jig appeared in 1900, followed by the Montgomery jig in 1910 and both have found extensive use. Increased interest in this field led to new designs, such as the American, Elmore; Burke and Faust. The first Baum-type Norton jig was operated in 1930 and within the last five years six large installations of this type have been made by the McNally-Pittsburg, Jeffrey and Link-Belt Companies. Wet-washing tables were introduced in 1922 but not extensively used until recently. The two most popular types are the Deister-Overstrom and Deister Plat-O. The coal-cleaning equipment used in Alabama has been relatively simple, yet efficient, but as the amount of impurities in the raw coal has been increased by mechanical mining and labor costs have become higher, much interest has been shown in sand cones, froth flotation and

Jan 1, 1945

By Robert F. Mehl

THE purpose of a symposium, I take it, is to gather together a group interested in a common intellectual field, in order to exchange views, and to appraise the state of knowledge. Symposia are the more successful, the greater the activity in the chosen field and the wider the range of applicability of the information gained. From time to time, as emphasis changes from one field of research to another, a given subject may enjoy popularity and later retire into a scientific somnolence, for fashions change even among engineers and scientists. The subject of this symposium, however, enjoys the advantages of wide applicability and of high current interest and productivity in results. It is a timely subject for a symposium; with much new information, surely metallurgical engineers should now devote an increasing attention to its application to practice. In beginning the symposium, I shall attempt briefly to appraise our basic knowledge in the field, to survey instances of practical usefulness, and to suggest where new work might be most productive of useful results. There are many examples of direct practical application of diffusion knowledge, and it is the purpose of the several papers in this symposium to review these. In this introduction I shall attempt to take a broad view, which, while it will but poorly serve the engineer in the solution of any given problem, may perhaps be of some use in assessing the state of knowledge in this field. RATE OF DIFFUSION Liquid Metals The process of diffusion among liquid metals, of gases in liquid or solid metals, or among solid metals, is ubiquitous, to say the least. The manufacture of alloys by the intermelting of metals, the addition of ferroalloys to molten steel, the deoxidation of molten alloys, all involve diffusion in the liquid state. The oxidation of molten copper for purposes of purification, the unsought oxidation of alloys generally during melting and casting, the reaction of liquid steel with oxidizing slags in the open-hearth furnace, the absorption of nitrogen by liquid steel, particularly in the bessemer furnace, all operate by the diffusion of gases in liquid metals. These processes have not been studied much from the viewpoint of rates of diffusion. Alloys are made by melting metals with stirring and with allowance of time for mixing (which is diffusion) but I know of no studies that seem to define the necessary times, and though in only very rare cases can faulty properties of the finished castings be ascribed to inadequate mixing, and though certainly in most cases, with stirring, the rate of diffusion is adequate to give sufficient mixing, a more quantitative treatment would be good to have. It is not wholly certain that thorough and sufficient mixing is at all times attained, for I suspect that ferroalloy additions to steel, and that additions of aluminum to steel as a deoxidizer, especially in the sampling of steels

Jan 1, 1944

By R. W. Gurry

IN the course of a series of measurements of the rate of diffusion of carbon in austenite at about 960°C. (1760°F.) and 1110°C. (2030°F.), it became necessary to determine carbon concentration when austenite is saturated with graphite, because of indications that the hitherto accepted value of the solubility of graphite in this temperature range is too low. Specimens of carbonyl iron were saturated at temperature in a hydrogen-toluene atmosphere which was in equilibrium with graphite. The carbon content of the steel was measured either by combustion or by the loss of weight after substantially complete decarburization in an atmosphere of hydrogen saturated at room temperature with water vapor. The absence of carbide or graphite as a phase at temperature was demonstrated by microscopic examination of saturated specimens that had been drastically quenched. The mean result of concordant measurements is that gamma iron saturated with respect to graphite contains 1.39 per cent carbon at 957°C. (1755°F.) and 1.89 per cent at 1110°C. (2030°°F.). These two values, plotted along with what appears from the literature to be the best value (0.69 per cent) at the iron-graphite eutectoid temperature (738°C. or 1360°F.), yield a curve that is almost linear, and which, continued upward through a short interval, leads to a limiting solubility of 1.98 per cent at the iron-graphite eutectic temperature1135°C. (2095°F.) instead of the commonly accepted value of about 1.7 per cent. EXPERIMENTAL PROCEDURE The samples of carbonyl iron, with the exception of those for microscopic examination, were cylinders 3.8 cm. long, 0.475 cm. in diameter. They were placed in a graphite holder and suspended in a vertical furnace, the temperature of which was controlled precisely by means of the resistance of the platinum winding.1 In the empty furnace the range of temperature over a distance of 4 cm, at the heat center was 9°C. (16°F.) but at a given location the temperature remained within ±0.5°C. (0.9°F.); it was measured by inserting a platinum-rhodium couple calibrated at the melting point of palladium, 1555°C. Since the presence of the sample probably improved the uniformity of temperature, it was decided to consider the average temperature of the sample to be 3°C. lower than the maximum measured in the empty furnace with the same setting of the controller. Data were obtained in the vicinity of two temperatures, 960°C. (1760°F.) and 1110°C. (2030°F.). The samples for microscopic examination, 2.2 cm. long, 0.475 cm. in diameter, were suspended individually on pure iron wire; for these no temperature correction was applied. The carburizing gas passed into the furnace was dry, oxygen-free hydrogen saturated at room temperature with

Jan 1, 1942

By Pierre J. Goossens

The coastal Cordillera around the Panama basin is made up of a Cu-rich tholeiitic volcanic belt representing either immature, late Mesozoic, island arc, or oceanic ridge. The same belt is exposed (uplifted) along the western slope of the Cordillera Occidental of Ecuador and Colombia. Ni-lateritic, Cr-podiform, and Pt-(Au) placer deposits are formed there after or within ultramafic bodies of oceanic association. Younger calc-alkaline igneous activity provided the mechanism for porphyry copper and associated ore formation. This Cu (+Mo) mineralization occurs along a narrow belt exposed by the erosion of the volcanic cap. Ag-Pb-Zn(-Au) and Hg volcanogenic mineralizations occur within the calc-alkaline belt, but slightly eastward and only above the copper belt. To the east and in the Venezuelan Andes, Paleozoic and older sediments, volcanics, and metamorphics contain volcanogenic massive sulfide, Pb-Zn, Au, and Cu sedimentary and vein deposits, and sedimentary iron ore and phosphate. The Cenozoic Cu, Ni-Pt-Cr and a portion of Ag-Pb-Zn, Hg, and Au mineralizations occur within or just above the oceanic basement whereas the eastern- most Cenozoic Ag-Pb-Zn mineralizations occur within and above a pre-Cenozoic sialic basement. This continental crust is the northern extension of the Peru-Chilean Andes where no oceanic crust has been preserved, however. Cenozoic metallogeny and subduction relationships are unclear in the Northern Andes. Yet strong structural volcanogenic ore controls exists apparently due to deep strike-slip movement between two plates.

Jan 1, 1977

By J. A. Herrick

Some time ago the writer needed a cheap, light, and portable hydraulic lift or crane, that would be universal in its application, and that might .be suspended from crane-arms, overhead tramways, beams, etc., and serve all the uses of ordinary hydraulic cranes. This was accomplished by the use of a closed cylinder of wroughtiron hydraulic pipe, provided at its top with the ordinary gland and stuffing-box, through the centre of which passed another similar pipe of slightly smaller diameter, forming in effect the ordinary plunger of the regular hydraulic crane. Over the top of this centre pipe or plunger or piston a disk of iron is placed with two arms projecting several inches from the rim, as may be required. Through these arms long iron rods pass, one on each side, and are fastened in place by nuts screwed to a bearing immediately above and below the disk. These rods also pass through the enlarged head of the lift down through a casting which closes the bottom of the cylinder, which, at the same time, gives them a bearing, and they terminate at a point below the casting equal to the height of the lift of the crane. The rods at this point pass through a disk similar to the top disk and are similarly secured in place. A common hook arranged with a swivel passes through the centre of the disk, and on this hook the load is suspended. For a stationary hanging crane the apparatus is suspended by two rods passing through the head of the lift in similar manner to the two already described, at right' angles to the first, and of sufficient length to let the first set play between them and rise to the full height which the plunger will allow. The second set of rods terminates in an iron disk similar to the one at the bottom of the first set, and a hook through this in turn suspends the entire apparatus. In order to adapt the lift to an overhead tramway, single or double, in place of the suspending rods, a projection or offset is arranged on each side of the head of the lift and a journal cast in, allowing the lift to be suspended on pivots arranged as follows:

Jan 1, 1879

By Leonard Larson

BEFORE direct or wet smelting of copper concentrate was adopted at the McGill smelter, in November 1932, actual furnace smelting tests had indicated the possibility of smelting between 400 and 500 dry tons of wet concentrate per furnace day. Under calcine smelting practice, solid charge up to 1165 tons per furnace day had been smelted. It was realized that in order to smelt wet tonnage in equivalent amount per furnace day, a higher combustion rate and higher temperatures would be neces-sary within the furnace. The reverberatory furnace considered in the following history of wet-smelting de-velopments is a unit 127 ft. long and 28 ft. 8 in. wide, inside dimensions (Fig. r). This furnace was in operation through all the years from 1934 to date. During 1933, the first complete year of direct or wet smelting, the plant was on a curtailed copper-production basis and the available concentrate tonnage to be smelted was limited. The average coal-firing rate in the reverberatory for that year was 124 tons per furnace day. This rate was lower than the average for years immediately preced-ing under calcine smelting practice. With this coal-firing rate, an average of 391 tons of new metal-bearing material (N.M.B.M.) was smelted per furnace day. The following year, 1934, the coal-firing rate was in-creased to 139 tons and the smelting rate increased to 448 tons N.M.B.M. per fur-nace day. In succeeding years production required that either more tonnage be smelted in the one furnace or an additional furnace be placed in operation. With this situation confronting the smelting depart-ment, it was determined to continue the program of increasing the coal-firing rate and also to intensify the temperature in the smelting zone of the furnace. In order to accomplish this, various improvements were made to aid combustion within the furnace. The burners were enlarged, in-creased primary air was supplied to the burners, and changes were made in the furnace to enlarge and streamline the out-let, thus increasing the furnace draft.

Jan 1, 1940

By N. S. Stoloff, M. Gensamer

Pyramidal (1122) slip was observed in cadmium single crystals deformed in compression and bending at room temperature and —196°C. Crystals tested in tension twinned with no evidence of pyramidal slip. A critical resolved shear stress law for the propagation of twin nuclei appears probable. Twinning and low-stress kinking are believed to inhibit the appearance of pyramidal slip under ordinary experimental conditions. SEVERAL reports of nonbasal slip in cadmium have appeared in the literature recently, two of the reports having appeared after work on this project was in progress. Brown1 reported cross slip on electron micrographs, but did not identify the cross-slip plane. Wernick and Thomas2 observed nonbasal slip traces parallel to [1123] on a crystal bent with a pair of tweezers, but this too was an isolated observation, with no confirmation of the glide plane. The critical stress for pyramidal glide, and the temperature and strain rate dependence of glide on this system were not discussed. price3 has reported some observations of nonbasal slip in thin films of cadmium with the aid of transmission electron microscopy, but a full publication of these observations was not available at this writing. Pyramidal (1122) <1123> slip also has been observed on zinc single crystals4, 5 and (1122) <1010> slip on magnesium single crystals,6 all oriented with the basal plane nearly parallel to the specimen axis. Since the choice of slip system in the hexa- gonal metals is generally believed to depend on the axial ratio, with basal slip predominating for high-axial ratios, the occurrence of pyramidal slip in zinc and cadmium is unexpected. The objectives of the present investigation were to confirm the existence of nonbasal slip in cadmium, to identify the slip system, and to obtain an estimate of the critical resolved shear stress, assuming that the resolved shear stress law is applicable. EXPERIMENTAL PROCEDURE Two purities of cadmium, 99.95 pet and 99.99+ pet, were employed. As received bars were cold swaged to 3/16 in. rounds. Bicrystals and tricrystals were prepared by passing the polycrystalline rods through a gradient without seeds. It was desired to obtain a large yield of single crystals oriented xo* < 5 deg, in order to be able to suppress basal slip during deformation by keeping the resolved shear stress on the basal plane to a minimum. Cylindrical single crystals of cadmium were prepared by a seeding technique developed by Weiner,7 to obtain a predetermined orientation. The single-crystal rods were sectioned to specimen size by an acid-cutting technique originally developed for zinc crystals.&apos; An aqueous solution of nitric acid was the cutting agent. The acid was picked up by a plexiglass disc spinning at an optimum cutting speed through the acid, which was contained in a sump-like arrangement. The section to be cut, the acid bath, and the spinning disc were enclosed in a plexiglass case. The rod to be sectioned was coated with a thick layer of collodion. At the points at which sectioning was to be effected, the collodion was removed over a length of about 1/8 in. by dissolving it in acetone. To prepare a surface suitable for microscopic exam-

Jan 1, 1962

By Irwin R. Higgins

CONTENTS A. Introduction B. System Description C. Uranium Pilot Plant Programs D. Summary INTRODUCTION The Chem-Seps Continuous Countercurrent Ion Exchange (CCIX) contactor was first conceived of in 1951 at the Oak Ridge National Laboratory. In the 27 years that have passed, there has been a steady stream of pilot plant process development studies on a variety of processes far too numerous to mention. These have been in the area of water and waste treatment, chemical processing, ore leach treatment, and many others. Many successful production systems have been installed and operated in a variety of applications. A few of these are described herein together with a description of the system and its operational characteristics. SYSTEM DESCRIPTION Chem-Seps Continuous Countercurrent Ion Exchange* is a single loop in which various sections are separated by valves to isolate the sections performing different functions simultaneously. (Refer to Figure 1). To explain the working of a typical unit, let us use a simple configuration used in water softening or demineralization. The resin is moved in small increments by the hydraulic ram action. All solution flows are countercurrent to the flow of resin. The contaminants are removed in the process cycle and the resin is moved in the pulse cycle. A controller in the effluent stream from the treatment section senses the proper limit permissible in the effluent. In the process cycle, the influent enters through a set of distributors, moves downward and leaves through a set of collectors. Simultaneously the regenerant enters through valve RE and leaves through valve WA. The rinse enters through valve RI and its duration is controlled by an interface controller. In the process cycle, valves B, C, and D are closed, isolating all the sections. The duration of the process cycle is 4-10 minutes depending on the contaminants to be removed and is controlled either by a timer if the contaminants are fixed, or a controller in the effluent stream, if the contaminants fluctuate at short intervals. Valve A is open and resin is dropped to the pulse section simultaneously with the removal of contaminants in the treatment section, regeneration of loaded resin in the strip section, and rinsing of regenerated resin in the rinse section. As the resin drops through the valve A, it is backwashed to remove the suspended solids or resin fines. Removal of these fine particles is important to avoid the unnecessary pressure drop in the bed and plugging of the distributors. At the end of process cycle, all flows going into the unit and out of unit are automatically shut off. The valve A closes and valves B, C, and D open in proper sequence. A slug of water enters through valve PU, a predetermined incremental volume of resin moves into the regeneration section, an equal volume of regenerated and rinsed resin enters the bottom of the treatment section, and an equal volume of the loaded resin leaves the top of the treatment section and goes into the resin reservoir section. (The resin movement is plug flow with the Reynold&apos;s number being in the laminar region of the curve, and no fluidization occurs during the pulse cycle.) The total time required for proper sequencing of valves and the movement of resin is 20-30 seconds. At the end of this period, the valves B, C, and D close and valve A opens. The solution valves open and normal steady state operation begins. All valves are interlocked and properly sequenced. The resin movement is precisely controlled by sonic devices. Special Features of Chem-Seps CIIX "Downflow"*: The raw feed in the treatment section moves downward and the resin moves upward in the treatment section. The "downflow" of feed overcomes the handicap encountered in "upflow" systems. The "upflow" of feed requires a certain minimum flux during process cycle to lift the bed and keep it

Jan 1, 1979