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By H. H. Stoek

(Wilkes-Barre Meeting, June, 1911.) THE origin of the Mine-Rescue Service of the State of Illinois can be traced to two distinct sources, the work of the Rescue Station at Urbana and the Cherry disaster. During the early part of the year 1909, the Technologic Branch of the U. S. Geological Survey, now the Bureau of Mines, in connection with the Illinois Geological Survey and the College of Engineering of the University of Illinois, established at the University of Illinois, in Urbana, a branch rescue-station to supplement the work of the Pittsburg station of the Geological Survey. As a result of the work of training at the station in Urbana by R. Y. Williams, mining engineer, and James Webb, foreman of the Bureau of Mines, and the use of the helmets at several mine-accidents in the State of Illinois, . the people of the State were somewhat familiar with oxygen-helmets when the Cherry disaster occurred, in November, 1910. The oxygen-helmets were successfully used in connection with that disaster, and upon the recommendation of the Illinois Mining Investigation Commission, the Legislature of the State, assembled in special session during the winter of 1910, passed a bill appropriating $75,000 for the erection and maintenance of three rescue-stations, stipulating that they should be located in the northern, central, and southern parts of the State. The Act also provided that the stations should be in charge of a Commission of seven, two representing the United Mine Workers of Illinois, two the mine-operators, one the Federal Bureau of Mines, one the State mine-inspectors, and one the Department of Mining Engineering of the University of Illinois. This Commission was called together by the Governor of the State, Aug. 2, 1910, and since that time three stations have been located, built, and equipped: at La Salle for the northern

Dec 1, 1911

By David R. Mitchell, R. M. Fleming

MANY trends were evident in the coal industry during 1951. Some were favorable for the industry; others were not. Probably those having the most far-reaching consequences are those affecting coal's markets. The rapid dieselization of many railroads and the continued trend toward oil and natural gas for home heating has caused serious dislocations in various mining districts. Mines have been forced to close down or to work on a reduced time basis. Conversely the increased activity in the metallurgical industries has increased the demand for coking and special purpose coals, and the growing production of electric energy has increased the demand in certain areas for steam coals. The increased demand for metallurgical coal and for steam raising by the large utilities is more than equal tonnage-wise to the loss of coal to railroad dieselization and conversions to oil and gas in the domestic market. Since much of the expansion is at captive mines of large corporations, many commercial operators supplying railroad and domestic users have had their markets curtailed without any compensating demand from other sources. Feverish activity was in evidence in many areas by commercial operators attempting to find new markets. The Norfolk and Western Railway is the, only major American railroad committed to the burning of coal. It has an exceptionally good cost record. During 1951, fifteen additional modern coal-burning steam switching locomotives were authorized for construction in the company's Roanoke shops at a cost of about $1,450,000. When the new order is completed the road will own a total of sixty locomotives of the modern-switcher type, including thirty acquired by purchase last year. The Norfolk and Western today is the only American railroad building its own motive power. All of these locomotives burn coal. Production Coal production in the year 1951 will approximate 570 million tons of anthracite, bituminous coal, and lignite. Of this, approximately 7 pet is anthracite. Total production is approximately 30 million tons more than in 1950 and reflects the increase in general industrial activity. Exports are expected to be close to 50 million tons. The industrial stockpile is close to 75 million tons, the highest it has been since 1943. With the exception of the usual local stoppages the year 1951 was unmarked by labor disputes. It was a year of good labor-management relations with a negotiated wage increase unmarred by a work stoppage. This may be the beginning. of an era of sensible negotiations to find agreeable settlements of disputes without the usual protracted bickering and industry-wide strike. The United Mine Workers of America has condemned wildcat strikes, stating that in nearly every, instance proper grievance procedures have not been instituted. Government defense organizations, perhaps profiting by their experiences during World War II, are pursuing a more enlightened program with regard to the fuel requirements of the nation. Many prominent officials have stated that wherever possible coal should be considered the primary source of fuel in all defense expansion. With the increased consumption of steel by defense projects, approval of transportation facilities for liquid and gaseous fuels have been refused where these fuels would enter a coal producing area. Liquid and gaseous fuel producers also have failed to provide satisfactory evidence that they have sufficient reserves and are capable of supplying their product in adequate amounts for the periods of time required. Realization has at last become prevalent that the hydro-electric power installations in several areas are insufficient to handle increased demands.' Steam generating plants are being constructed in these areas and equipped to burn coal. Most notable of these, is the TWA area whose supplemental steam power plants are expected to consume 10 million tons of coal per year when completed. The Pacific Northwest is another area which is expected to follow the same pattern. Nearly all new steam-power generating plants are being equipped to burn coal either as a primary fuel or as an emergency measure. Metallurgical Coal The steel and other metallurgical industries are expanding rapidly to meet increased industrial demands and defense orders. Since these industries are dependent on coke, coke has had a year of high production and can be expected to break all existing records when steel-plant expansions are completed. Most notable of the expansion projects is that of the Delaware River area where the iron ore will be received by ship from Canada and South America. U. S. Steel is only one of several steel corporations that are planning on putting integrated steel producing units in this area. The Defense Solids Fuels Administration announced in December that through. November 30, 1951. 29 projects in eight states, with a total cost of

Jan 1, 1952

By Arthur L. Walker

In June, 1893, while I was with the Old Dominion Copper Co. at Globe, Ariz., I received an offer to take charge, as manager, of the Baltimore Electric? Refining Co. which had recently built a plant at Baltimore, Md., having a capacity of about 1200 tons of electrolytic copper per month, using the Hayden series system. This was a very large output, for in those days the plants were small and the art was in its infancy. At the beginning of 1893 there were 11 electrolytic refineries in the United States, most of them using the multiple system. The Hayden series system was used at Baltimore and Bridgeport and the Smith series system was in operation at the Electrolytic Copper Co. at Ansonia, COM., and at the Pennsylvania Salt Co. at Philadelphia. In the latter system the plates were placed in a horizontal position, but the operating difficulties were so numerous it was soon discarded. The total production of these 11 plants in 1892 was very small; I have not been able to discover reliable figures of production. There were about thirty electrolytic copper refineries in the world that year, according to Titus Ulke, and the entire production of these

Jan 1, 1932

By Kim C. Harden

INTRODUCTION The purpose of the following discussion is to present the state of the art of solution mining. Since the economics of a mining method ultimately determines its applicability and viability this presentation shall revolve around the costs of in- situ solution mining. First the assumed physical characteristics of the hypothetical ore body are described, followed by the appropriate operating assumptions. Then after a brief discussion on the type of surface plant to be used, the assumed project time tables and costs for Texas and Wyoming are presented. Finally, the economics of in-situ uranium leaching are analyzed through the use of discounted cash flow rate of return analysis. ORE BODY CHARACTERISTICS The assumption of the ore body characteristics is probably the most variable portion of this discussion. The characteristics that have been used are based mainly on state of the art technology, however, consideration of the most common depths of ore, ore thicknesses, and permeabilities also influenced these assumptions. In addition, it is assumed that these assumptions are equally applicable to Texas and Wyoming. The average grade of the ore is assumed to be .09% U308 with no apparent disequilibrium. The average thickness of ore is 2.29 m (7.5 ft) which results in an average grade-thickness (GT) of .675. The assumed depth to the top of the ore is 121.92 m (400 ft), the ore density is placed at 1.78 gm/cc (18 cu ft/ton), the porosity is estimated to be 28% and the permeability 1 darcy. These assumed ore body characteristics are shown in Table I. In addition, it is specified that the costs to be later discussed are based on a minimum GT cut-off of 0.15. It is more common to use GT cut-offs of 0.30 to 0.50 but GT cut-offs as low as 0.15 in conjunction with a minimum grade of 0.05% U308 have been used in the past with success and is considered state of the art. The ultimate percentage of uranium recovered from the ore is left to the discretion of the reader since the costs and economics are based on pounds recovered by the surface plant. OPERATING AS.SUMPTIONS An annual production rate of 200,000 lbs U308!yr was chosen for this example. In order to maintain this production rate, based on the ore body characterized above, a flow of 4731 liter/min (1250 GPM) with a recovery solution grade averaging .039 gm U308/liter is assumed. A regular 5 spot well field pattern is used with a well spacing of 21.5 m (70.7 ft) between like wells and 15.24 m (50 ft) between unlike wells. This well spacing gives each well an area of influence equal to 232.25 sq m (2500 sq ftl. An excess wells factor of 1.17 is used to estimate additional monitor wells and well field boundary wells. Each production well is expected to yield an average flow rate of 37.85 liter/min (10 GPM). In addition it is assumed that the ore body has a good shape in that it is not tenuous and narrow but has at least an average width of 200 ft. The process chemistry required for this ore body is assumed to be based on the sodium carbonate System- Oxygen is the chosen oxidant. Sodium chloride elution followed by precipitation with hydrogen peroxide makes up the remaining portion of the process. A fluidized up-flow ion exchange system is specified. The operating assumptions are listed in Table II. Restoration of the ore body shall be assumed to be accomplished through the use of ground water flush. Other methods may be considered as having to fall within the costs estimated for a ground water flush in order to be economic. In Texas it is assumed that a high capacity disposal well (200 GPM +I is required and in Wyoming evaporation ponds covering approximately 35 acres are to be used. No specific cost has been given to restoration. Instead only the additional capital investment for restoration equipment is given. Then, one year of restoration operating expense is estimated and included as the operating expense for one year beyond the last pound of U308 produced. It is also assumed that restoration will be pursued in the mined out areas of the ore body contiguous with ongoing production.

Jan 1, 1980

By John S. Alexander

DURING the past few years the block coal of Indiana has been talked about and written upon to such an extent, that almost every one at all interested in such subjects, has been made acquainted with the area and geographical position of the field, the quality and peculiar properties of the coal, the thickness of the veins, the manner of mining, etc. The writer, therefore, taking for granted that, to those present all this is familiar ground, desires more especially to call attention to the position this coal takes in the market, and at those points of consumption where it is brought in direct competition with rival fuels. Chicago, thus far, has been the chief consumer of Indiana block coal, the distance from the mines at Brazil, via Terre Haute and the

By John S. Alexander

DURING the past few years the block coal of Indiana has been talked about and written upon to such an extent, that almost every one at all interested in such subjects, has been made acquainted with the area and geographical position of the field, the quality and peculiar properties of the coal, the thickness of the veins, the manner of mining, etc. The writer, therefore, taking for granted that to those present all this is familiar ground, desires more especially to call attention to the position this coal takes in the market, and at those points of consumption where it is brought in direct competition with rival fuels. Chicago, thus far, has been the chief consumer of Indiana block coal, the distance from the mines at Brazil, via Terre Haute and the 25

Jan 1, 1873

By Charles R. Keyes

I. INTRODUCTORY. IN the dry regions of the globe many silver-deposits display certain remarkable features which at the same time are so totally unlike anything met with among ore-bodies elsewhere, that they have long presented exceptional difficulties not only to a satisfactory explanation of their origin, but to economical milling. A most striking peculiarity of these ores is the haloid character of many of their bonanzas. To this fact more than to any other may be ascribed most of the unusual uncertainties of mining-operations in desert countries-the marvelous local richness of the ores under certain conditions and the often abrupt transition, to utter poverty; the inadequacy of ordinary prospecting customs and the constant devising of new exploratory methods; the unaccountable losses in smelting and the continual change in treatment-practice. For these anomalies certain climatic conditions seem to be chiefly responsible.

Jul 1, 1911

By Herbert H. Kellogg

The chemistry of sulfide roasting is analyzed to show those aspects of performance which Thecan be predicted from considerations of thermodynamic equilibrium. It is concluded that equilibrium calculations can serve as a yardstick for prediction of roaster-gas composition and sulfate formation in the calcine for fluid-bed roasting practice. A method of calculation which combines the equilibrium and stoichiometric factors is outlined, and illustrated by an example from the roasting of Cu-Fe sulfides. ON first examination the roasting of sulfides would not seem to be a process to which equilibrium calculations could be applied to yield results of practical value. The burning of sulfides is an ex-tremely spontaneous reaction, excess air is always used to obtain rapid reaction, and in most traditional roaster designs (hearth roasting) the temperature and gas composition vary widely from point to point in the roaster. These are all conditions which promote a non-equilibrium state in the roasting process. However, the introduction of the fluid-bed roaster1 has brought to roasting practice new conditions of operation, so that equilibrium calculations can be used to predict certain limited aspects of practical operation. The fluid bed is characterized by uniformity of temperature, gas composition, and calcine composition from point to point throughout the bed. These relatively uniform properties of the bed make the meaningful calculation of the equilibrium state of certain reactions in roasting possible. It is the purpose of this paper to analyze roasting from an equilibrium viewpoint, to show where the method may be expected to yield valuable information, where it fails, and to outline methods of calculation. Main Roasting Reaction The reaction MS (s) + 3/2 0, = MO (s) + SO, [I] might well be called the main roasting reaction. It adequately accounts for the well known facts that metal sulfide and air are the main reactants, and metal oxide and SO, the main products. Deviations from this simple stoichiometry will be discussed under side reactions. For all metal sulfides, Reaction 1 is strongly exothermic and the equilibrium position is far to the right. Since the reaction is exothermic, the equilibrium shifts in the right-to-left direction as the temperature is raised. However, at all practical roaster temperatures (between 500" and 1200°C), the amount of this shift is small, and the equilibrium position remains far to the right. The main roasting reaction is essentially irreversible, and the application of equilibrium considerations to this reaction will not be! a fruitful field of investigation. A knowledge of the rate of this reaction. although not the proper subject of this paper, is of great importance to the roasting process. Analysis of the factors which influence this rate shows that: 1) The rate will increase rapidly with increased temperature. 2) The rate will be proportional to the partial pressure of oxygen at the surface of the reacting particle. Because of diffusion effects, the partial pressure of oxygen at the particle surface may be considerably less than in the bulk of the roaster gas. For this reason, intimate contact between gas and solid and high relative velocity between gas and solid, such as can be obtained in a fluid bed or flash roaster, will markedly increase the rate of reaction. 3) The rate will be proportional to the surface area of metal sulfide. Therefore, smaller particle size will permit a more rapid rate per unit of weight. To achieve practical rates of roasting the partial pressure of oxygen in the roaster gases must be maintained at a finite level at all times. Even where the gas-solid contact is very good, as in the fluid-bed roaster, the oxygen content of the gas must not be allowed to drop below some limiting value (perhaps 1 pct), or the rate of the main roasting reaction will become too small for practical operation. Side Reactions in the Gas Phase The gas phase during roasting of metallic sulfides contains the elements oxygen, sulfur, nitrogen, by-

Jan 1, 1957

By Frank Firmstone

It is proposed to consider here only comparisons made between results obtained when the materials employed are precisely the same, two furnaces at the same works for example, or the same furnace under different conditions as respects temperature of blast, method of charging, etc. With this limitation, the most important point in the comparison is the amount of fuel consumed per ton of iron. This, in fact, is the only item in the prime cost of the iron which can properly be said to vary in consequence of changes in the furnace process; for of the other principal items, the ore used per ton of iron, and also the amount of fluxing material, will depend on its richness, which, under the assumed conditions, will be practically invariable when taken over any considerable period; the labor cost also, from its nature, should be very nearly constant when estimated on the ton of iron. Moreover, as it consists principally in handling and moving material, it is greatly affected by circumstances quite independent of

The Climax Molybdenum Company' s newest operation is the Henderson project, which is located on the east side of the Continental Divide. Ore reserves are estimated at more than 300 million tons at a grade of 0.49% MoS2. For mining the ore a modified panel caving system is used. The ore is transported to the Henderson mill on the west side of the Continental Divide on a 42 in. (1.07 m) gauge railroad through 9.6 miles (15.4 km) of tunnel and 4. 8 miles (7.7 km) of surface track, a total distance of 14.4 miles (23. 2 km). Each train consists of 30 ore cars with capacity of 22 st (20 mt) of ore or 660 st (599 mt) per train. The unmanned trains operate completely automatically. At the 54 in. by 84 in. primary gyratory crusher, the bottom dump ore cars pass over a dump scroll to discharge at a maximum rate of 100 st (91 mt) per minute. Dumped ore falls onto a rock box before passing over a 17 ft long by 46 ft wide (5. 2 m by 14 m) grizzly set at a 40° angle with bar spacings of 7 ½ and 8 ½ in. (191 mm and 216 mm) openings. An estimated 30 to 50% of the ore (which at times is wet and sticky) passes through the grizzly. The operators report that a shallower angle on the grizzly, perhaps 35º to 37º, would have been sufficient and would have prevented impact on the back of the crusher pocket by large boulders. This pocket as well as the crushed ore pocket under the crusher is lined with AR plate. The accompanying drawings show a "future splitter" and the operators are planning the installation of this unit to prevent impact and subsequent damage to the crusher spider. Crushed ore at an 8 to 9 in. (203 mm to 229 mm) size joins the grizzly undersize in the pocket beneath the crusher. This pocket is sized at 660 live st (599 mt) so that it will accept an entire trainload of ore. Crushed ore is removed from the pocket by a four-speed, 84 in. (2. 1 m) apron feeder operating at 9.6 to 29. 3 fpm (2.9 to 8.9 mpm). A 60 in. (1.5 m) takeaway conveyor operating at 480 fpm (146 mpm) delivers the crushed ore to a 58,000 st live and 193,000 st total (52,600 and 175, 100 mt) open stockpile under an A -frame supported tripper gallery. Impact idlers on the takeaway conveyor under the apron feeder discharge are mounted on wood blocks for a cushioning effect and ripped belt detecting bars are located behind each impact idler. The conveyor tunnel is of corrugated and galvanized steel construction rather than concrete. The concrete crusher enclosure is circular in design from its lowest level at elevation minus 110 ft 4 in. (minus 33. 6 m) to elevation

Jan 1, 1978

THE research and policy committee of the Committee for Economic Development, a non-profit research organization composed of leaders in industry and the professions, including such prominent figures as Dwight D. Eisenhower, Paul G. Hoffman, and Charles E. Wilson, recently published a report entitled "How to Raise Real Wages." Since its founding in 1942 the CED has initiated many objective studies on matters of national economic policy, and has gained a fair measure of prestige from its activities. The CED's latest report cites the more than three- fold rise in real wages in the United States in the past 50 years and attributes the accompanying improvement in the standard of living to the great in- crease in production per man hour. The report states that in order for production to continue rising in the future with the attendant increase in real wages four basic conditions must be manifest: (1) better methods, (2) more capital, (3) better training of workers, and (4) better management. The report cautions that although prospects for increases in output per man hour are good "it is desirable that steps be taken to encourage such a rise."

Jan 8, 1950

By H. C. Otis

DURING the past year or two considerable time and money have been spent in developing equipment for tubing large-volume high-pressure oil and gas wells without loss of production. That the efforts have met with success is proved by the fact that any properly cased well, regardless of its production or regardless of its rock pressure or working pressure, can now be safely tubed with no appreciable loss of production, provided experienced crews handle the new equipment. It is conceded that a far better gas-oil ratio can be secured by flowing a well through tubing of a proper diameter-the diameter of the tubing depending, of course, upon the volume of fluid and gas and the depth of the well-than by the more usual method of installing a choke at the mouth of the well. Where possible it is becoming the practice to bring wells in through tubing, though in the majority of instances this is not feasible. In many instances it has developed, after a well has been brought in through tubing, that the tubing installed is either too large or too small for the most efficient operation or that the well had not been drilled deep enough. The equipment described in this paper, in the hands of experienced men, permits the tubing of these wells after they have been brought in, with the size of tubing then determined by actual observation to be best suited for each particular well, and with no danger of the well getting out of control, of catching fire, of drilling itself deeper, or of any loss of production. The deeper the flowing production with the consequent future higher lifting costs, the more essential it is that the size of tubing used be such as to give the best possible gas-oil ratio. The ability to bring in these deep wells through the casing and then, immediately after their completion, regardless of volume or pressure, to tube them with the size of tubing that permits the most efficient use of the gas present will doubtless save many dollars.

Jan 1, 1929

By N. S. Stoloff, R. C. Ku, R. G. Davies

The stress-strain behavior of Fe-Co and Fe- V alloys containing up to 25 pct solute have been studied in the temperature range 25° to - 196°C. The microyield stress is independent of temperature for all alloys studied, while the temperature dependence of the ordinary yield stress is large and comparable to that for unalloyed iron. An alloy-softening phenomenon is observed for many alloys relative to unalloyed iron, and appears to be a consequence of removal of inter-stitials from solution and their subsequent agglomeration. The ductile to brittle transition temperature is increased markedly by both cobalt and vanadium. Transmission electron microscopy observations reveal that cross slip and cell formation are restricted with increasing solute, which can account for many aspects of yielding and fracture behavior. IRON and other metals of bcc structure slip on many planes of the (111) zone at moderate and elevated temperatures, giving rise to wavy glide. At low temperatures, however, there is a tendency for slip to be restricted to planes of only one or two crystallographic types; in the case of Fe-Si alloys for example, (110) glide predominates, giving rise to planar slip bands."' Restriction of cross slip at low temperatures in iron has been linked by Brown and Ekvall3 to increased strain hardening in the preyield microstrain region, leading to an apparently large temperature dependence of the yield stress. Support for this view has been provided by Lawley and Gaigher4 from a transmission electron microscope study of molybdenum. They concluded that dislocations at lower temperatures, by virtue of their inability to change planes readily, are unable to interact and annihilate, leading to an increase in internal stress and work-hardening capacity. Restriction of cross slip at low temperatures has also been suggested to be an important factor in determining the ductile to brittle transition of single-phase solids, including metals of bcc structure.5 It is pertinent to note that most solute elements tend to raise the ductile to brittle transition temperature of poly crystalline ferrite,6,7 although in some cases, e.g., Fe-Cr7 and Fe-Al,8 the dilute alloys may actually yield at lower stresses than unalloyed iron. It appears, therefore, that solute elements may have an effect analogous to lowering the test temperature with regard to restricting cross glide. Such an effect has indeed been observed in the ionic alloy system KCl-KBr,9 in which bromine atoms, although causing little misfit and consequently little effect on the yield stress of KC1, have a pronounced effect in restricting wavy glide and inducing brittleness. The purpose of the present investigation was to determine the effects of selected solute atoms on the yield stress and fracture behavior of iron, and to elucidate in turn how the slip character was related to these properties. Since it is known that interstitial atoms exert a strong influence on the mechanical behavior of iron, the solute elements chosen for study were vanadium, which reacts strong with interstitials,10 and cobalt, which does not form highly stable interstitial compounds." EXPERIMENTAL PROCEDURE Four-pound ingots of each of the following alloys were vacuum-melted in zirconia crucibles: Fe-1, 5, 10, 25 at. pct Co, and Fe-1, 4, 10, 20 at. pct V. An ingot of unalloyed iron from the same original stock as that used for the alloy was remelted under similar conditions to eliminate variations in interstitial content arising from differing thermal histories. Chemical analyses of the iron, cobalt, and vanadium melt stock are listed in Table I. All ingots were rolled and swaged, commencing at 850°C, to 1/4-in.-diameter bar. The bars were then annealed at 850°C for 1 hr (except that the 10 and 20 pct V alloys were heated for 2 hr) to give a uniform equiaxed grain size of about 0.05 mm. Tensile samples of 0.125 in. diameter by 0.8 in. gage length were utilized for studies of the ductile to brittle transition. A second set of tensile samples were utilized for microstrain experiments in which strain sensitivity of 1 x 10"6 was achieved by a capacitance method.3 Cylindrical compression samples 0.250 in. diameter by 0.4 in. long also were prepared to study the influence of composition on the macroscopic yield stress, and for metallographic analyses of defor-

Jan 1, 1965

By Moore, Richard B

PROBABLY no other metal excites as much interest, among both scientific men and the general public, as radium. This is due partly to the high cost of radium salts and partly to the peculiar properties of the element. Since radium-bearing ores were discovered in the United States, the interest of American scientific men has been stimulated and, at the present time, more radium is extracted and refined in this country than in all the rest of the world together. HISTORY The property of radioactivity was discovered, partly by accident, by Henri Becquerel, the French physicist, in 1896. He was experimenting with certain fluorescent substances in order to find, if possible, a connection between fluorescence and the recently discovered Y-rays. Among other chemicals which possess the property of fluorescence, he was using

Jan 8, 1918

By L. Adler

Two basic cases involved in the design of an opening in bedded rock are: 1) where the beds deflect from each other so as to be separated; and 2) where the beds deflect onto their lower neighbor, loading the latter while reducing their own load. Analysis of case 1 is relatively simple, while the interference of the original deflection curves in case 2 complicates analysis of the latter. Figures and formulae are presented for determination of interference loading in this situation. The design of an opening in bedded rock involves two basic alternatives.1'2 In one case the beds deflect away from each other, so as to be separated from their neighbors; in the other case the beds deflect onto their lower neighbor, loading it, while reducing their own load. This is shown in Fig. 1. The first case lends itself to a simple analysis, since the upper and lower boundaries of each beam remain a free surface. In the latter case, however, due to the interference of original deflection curves, the analysis is more complex. Earlier investigators of this interference problem solved it by making the tacit assumption that this type of loading was uniformly distributed along the beam's length. While there is no intuitive justification for this assumption, its use does lead to an approximate solution. Still, a more exact solution does not require this apparently unrealistic restriction. To achieve initial simplicity, the analysis will at first be limited to a two-bedded sequence, as in Fig. 2. Since both beams have identical end conditions and spans, the original deflections of their center lines are members of the same family of curves, as shown in Fig. 3. Therefore, if interference occurs at one point along their center lines, it will occur at every point. For convenience, the midspan point is that usually examined. Deflection at midspan for a uniformly loaded beam regardless of end conditions is: where c is a constant depending on end conditions; w is the body load and equals yh, lb per ft; y is the specific weight of rock, lb per ft'; L is the span, ft; E is Young's Modulus, lb per ft2; I is moment of inertia (for a rectangular section), I = bh3/12 = h3/12, (ft4); b is the width of beam and equals one; and h is the thickness of beam, ft; or 3U > yB an Substituting the appropriate values into Eq. lb, interference between the two beds can be predicted. Should it occur, the additional loads along the lower boundary of the upper bed and the upper boundary of the lower bed must be determined. From the principle of action and reaction, these added loads are known to be equal and opposite, as indicated in Fig. 4. Since the final deflection curves for both beds are identical (Fig. 3), their successive derivatives of y with respect tox must also be identical. That is:

Jan 1, 1961

By K. Prins

ONE of the more pressing problems faced by the coal industry today is the development of adequate means for meeting conservation laws, particularly those involving stream pollution, in various parts of the country. Discharge of dirty coal washery water into streams and rivers is almost universally frowned upon. Many states have enacted laws carrying heavy fines to curb the practice. The Prins stream-cleaner is one of the latest machines to enter the market. It is closely related to the cyclone thickener in principle. Eleven stream-cleaners are currently operating, ranging in size from 4 to 16 in. diam. In more recent installations the water enters directly in line and on a tangent with the impeller. The impeller consists of a vertical shaft up through a packing gland and bearings, and a V-belt pulley. The lower part is a tubing fastened to the shaft above, extending through the water intake compartment and provided with six vertical flat bars welded to the tubing. Portholes are situated in the upper end of the tubing, immediately below the point where shaft and tubing join together. The portholes are placed so that they are in open communication with the upper compartment of the stream-cleaner from which the processed water is discharged. The impeller is motor driven with a wide range vari-pitch drive employed between motor and impeller. The motor is mounted vertically, and the mounting provided with a vertical hinge allowing for needed adjustment of the wide range vari-pitch drive. The dirty coal washery water entering the machine under 20 lb psi pressure, flows from the compartment above the impeller, between the impeller blades, and is whirled around in the vertical section of the impeller enclosure. The speed of the impeller supplies centrifugal force and velocity required for separating suspended solids from water. The lower part of the machine consists of a cone, whose action is similar to other machines of the same type. The underflow discharge orifice is a cold rolled steel block machined to correspond with the cone angle and allows insertion of steel tubes of different diam. On 16 in. machines a 1 ¾ in. ID vertical discharge pipe is used. Provision is made for attaching a curved section of the same diam to the vertical pipe, to which, in turn, different lengths of horizontal pipe can be connected. Curved Pipe Advantageous It has been found that a curved pipe offers resistance to discharged material flow. In addition, the rotary motion of the underflow can be easily arrested in a curved pipe. Impeller speed of the 16 in. diam machine is provided from 400 to 800 rpm. A speed of about 474 rpm is suitable for maintenance of a constant underflow in coal slurry. In one installation 5x ¼ in. coal is cleaned in a Jeffrey Baum type washer at a 225 tons per hr rate. Washer installation is of the conventional type and a drag type sludge tank is used for water clarification purposes. Capacity of the water system, including the Baum washer, is about 40,000 gal. Before placing the stream-cleaner in operation, it was necessary to flush out the entire system every five days of two shift operations. The only time the system is drained now is for repair work on the sludge conveyor or the rig. The suction line of the stream-cleaner pump terminates in a number of small branch lines located at a depth of about 4 ft above the sludge conveyor. Each branch line extends the full width of the tank and is provided with four intake ports, each one with a funnel shaped inlet projecting downward. The arrangement provides an extensive pick-up area, for dirty water, and the inlets are arranged for a low rim velocity, preventing the taking in of coarse particles. The funnels are also arranged to extend up or down in the tank. They are set to pick up -60 mesh material exclusively.' The material is a high ash and high sulphur product and thus has to be disposed in the refuse conveyor. The underflow of the stream-cleaner is discharged on top of the washery refuse which is carried in a drag type horizontal conveyor, discharging into another refuse conveyor inclined at 30º with a short horizontal loading section. Some Disadvantages The impeller inherits certain disadvantages because of the nature of its construction. Additional moving parts make it subject to wear and maintenance costs. The advantage of being able to maintain a constant speed, however, to produce desired water velocity in the machine outweighs the drawbacks. Better separation between water and solids can be obtained by regulating time of residence of water through adjustment of valves in the intake and discharge lines. The amount of fines encountered during plant operation will vary because of higher or lower moisture in coal passing over fine coal vibrating screens. Even the amount of fines picked up by underground loading machines will be inconstant. Consequently, the percentage of solids will vary in water to be processed. The velocity in the feedlines to the slurry thickeners will fluctuate, with the required water velocity lacking. Another advantage advanced for the machine is its ability to operate on 15 to 25 lb line pressures at the water intake, reducing pump power required and pump maintenance.

Jan 1, 1952

By W. M. Mobley

The coal industry in Alabama, centered in Birmingham, has pioneered coal-bene-ficiation 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,' 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 re- ceiving 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 Lurig 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-Pitts-burg, 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, 1947

By W. M. Mobley

The coal industry in Alabama, centered in Birmingham, has pioneered coal-bene-ficiation 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,' 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 re- ceiving 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 Lurig 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-Pitts-burg, 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, 1947

By Henry P. Ehrlinger

The Alta and St. Louis mines of the Alta Mines, Inc., produce a somewhat oxidized ore with a talc gangue that presents quite a problem in milling. For several years the mill recoveries were relatively low and there seemed to be no cure for this condition even with increasing the various collecting agents and frothers to two or three times of what they should be. The usual alibi of "oxidized ore" was given as the reason for these results and the matter stood there for some time until these investigations were started. Flowsheet As a brief outline of the flowsheet the ore is crushed to minus I in. in a Telsmith jaw crusher and then fed to a 64½ Marcy grate mill. The resulting product from the grate mill is passed over a trommel where the coarse particles (plus 7 mesh) are returned to the grinding circuit through the classifier. The under 7-mesh pulp is passed over two Denver mineral jigs where a high grade concentrate is produced, thence over two Wilfley tables where a table concentrate is produced and about 80 pct of the oxide lead recovered. The tailings from these tables are pumped back to the classifier and after classification the 30 pct pulp density product, which overflows the lip, is passed into a series of 10 Denver economy flotation cells where a lead-copper concentrate is produced. Three products are made and shipped to the smelter termed jig, table and float concentrates. Results with various Reagents In March of this year work was started on the metallurgical problems of the mill and, as is shown in Table I, the results obtained were not too efficient. Tests were run with various reagent combinations that were recommended for ores of this nature but there was no appreciable effect on the recoveries. The grades of the flotation concentrates were reduced by the insoluble picked up and retained by these more powerful collectors and frothers. Results with "Metso" In past work on ores of this type the use of "Metso" had improved gold and lead recoveries, but no effect on the silver and copper recoveries had been noted. To

Jan 1, 1949

By Robert B. Hall, Charles W. Bauer

Alunite, KA13(SO4)2(OH)6, was used from the 15th century until early in the present century as a source of potash alum and aluminum sulfate. Since early in this century its use for this purpose has declined to virtually nil with acid treatment of bauxite or clay replacing alunite as a source of alum. During the First World War alunite served as an emergency source of potassium sulfate fertilizer in the United States and Australia. More recently, alunite has been investigated as a nonbauxite ore of aluminum, with potassium sulfate and sulfuric acid as recoverable byproducts. The Soviet Union established a commercial-scale plant in the mid-1960s producing cell-grade alumina from alunitic ore and recovering byproduct potassium sulfate and sulfuric acid. Interest in nonbauxite ores, including alunite, is expected to continue in countries that presently are dependent on imports of bauxite or alumina to support their aluminum industries. The US has the largest aluminum producing and fabricating industry in the world but imports more than 90% of its aluminum-bearing raw materials. This chapter reviews the status of alunite as a potential source of aluminum with byproduct fertilizer, and the resources available to support this industry, whenever this becomes economically feasible. Historical Background Alunite has received little attention until recently, yet for centuries it held a higher place than many of the industrial minerals considered essential today. The deposits of alunite at Tolfa, northwest of Rome, were mined almost continuously from 1462 until about the middle of the present century (DeLaunay, 1907; Lombardi, 1977). The potash alum made from the alunite, called Roman alum, was a valued commodity in international trade, being used in the making of paper, in tanning leather, and as a mordant in textile dyeing. The industry at Tolfa is said to have been controlled by the Vatican and was a source of revenue for the Papal State until fairly recent times (Lombardi et al., 1977). Since early in the present century, the preferred method for making alum has been by acid treatment of bauxite, clay, or other aluminous material, so that alunite has fallen into near-total disuse as a source of alum chemicals. Appreciable quantities of mixed alunite-kaolin rock still are quarried in the Tolfa district for use by the Italian cement industry (Lombardi and Mattias, 1979), but the role of this material in the making of Italian cement is not clear. A large deposit of alunite-bearing rock in Chekiang province, mainland China, is said to have yielded a "great quantity" of alum over a period of several centuries (Yih, 1931). During the 19th century, alunite deposits in France, Spain, Australia, and other countries were exploited as raw material for making potash alum and aluminum sulfate. Interest in alunite was revived during the First World War, not for alum, but as an emergency source of potassium sulfate fertilizer. Large veins of nearly pure alunite had just been discovered on Alunite Ridge, 11 km southwest of Marysvale in Piute County, Utah (Butler and Gale, 1912), and these were quickly brought into production for making potash fertilizer to replace supplies cut off from the famous deposits at Stassfurt, Germany. An estimated 225 kt of Marysvale alunite was mined and processed at that time (Callaghan, 1973). Similarly, Australian alunite deposits were mined as a source of potash fertilizer during both world wars (Hall, 1978).

Jan 1, 1983