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THE slag performs two useful functions in open-hearth steel- making. First, it is the means of disposal of all the impurities, save carbon, which are removed from the charge materials in refining the steel, so it must be suited to receiving and retaining as much as possible of undesirable elements with minimum loss of iron and useful alloying elements. Second, the slag serves as a buffer between the flame and the bath and transfers oxygen to the metal at a measured pace contrasted with the violent and difficultly controlled oxidation of the bessemer process. At the same time, the slag is a thermal insulator that limits the rate at which heat can be delivered to the metal from the flame, and a heavy slag blanket thereby serves as a brake on production rates. It has been said with justification that the secret of good steel- making is to make a good slag and the steel will take care of itself. In a broad sense, the subject of slag control includes not only the adjustment of the composition of the slag but also of its relative weight in terms of percentage of the metal-bath weight. The slag is made up essentially of: (1) oxides formed by oxidation of silicon, manganese, phosphorus, and iron in the charge, (2) Si02, MnO, MgO, and other oxides charged as such or introduced incidentally with ores or fluxes or by refractory erosion, and (3) lime charged or fed to the heat to neutralize the "acidic" oxides present (Si02 and P205). Obviously, if practically perfect control were possible over the composition of the charge as well as over the oxidation process during the melting period, both slag weight and composition would be very nearly predetermined; and the control problem would be solved. Unfortunately, only a somewhat imperfect control over these factors is possible, so the operator must begin during the lime boil and continue through the refining period to discover and correct departures from the normal pattern of slag development. To do this as the heat progresses, the first helper must be able to make quick estimates of the composition and physical state of develop-

Jan 1, 1964

By A. W. Allen

THE technological study of the treatment of gold and silver ores has been largely responsible for the phenomenal strides which have marked the progress in this branch of metallurgy during recent years. In no other application of science to industry is system more imperative at every stage; correct formulation of result is only second in importance to efficiency of operation. Progresss in method has been largely the result of individual initiative, aided by the work of technical journals and metallurgical societies. There still remains a field for profitable discussion, and it is my intention to deal with some inconsistencies of expression common to a number of writers on the subject; and also to draw attention to the false impressions created by a statement which may, possibly, be only unintentionally misleading. I also take the opportunity of tracing the source of both metallurgical and economic result; and to question the desirability of the methods usually employed in arriving at a final statement. In the first place, I wish to deal with the question of extraction. This term may mean anything or nothing, and it is often used to mislead. In a report of a mining engineer as to the result of operations subsequent to the remodeling of a treatment plant prominence was given to the statement that the extraction had amounted to 91 per cent. On examination it was found that only a small percentage of the so-called extraction was calculated on actual bullion return, a proportion being figured from the gold left in the zinc boxes, and not recovered. In addition to this, a proportion of the extraction was said to be due to gold in concentrate. As there was no concentrating apparatus on the property and as concentration did not form a part of the scheme of treatment the matter was investigated, with the result that the concentrate, of lower gold content that the original ore, was found to be the result of an inefficient pumping system which failed to provide for the whole of the battery product delivered to it. The balance was stacked and, although itself of no marketable value, the gold content was allowed to form a basis of extraction figures.

Jan 7, 1914

By W. N. Best

Crude oil attracted attention because of its excellence as a fuel for openhearth furnaoes; for making crucible steel and brass; for melting copper, lead, tin, zinc, nickel, silver, malleable iron, gray iron; and for the production of steam in all sizes and types of boilers. For heating furnaces, mold-drying ovens, core ovens, ore roasters, calciners, hotair furnaces, sand drying, asphalt road work, etc., its superiority over other fuels is due to the fact that the heat is at all times under perfect control, so that a constant temperature may be attained and maintained at the will of the operator. This mcans a maximum amount and uniform quality of work. Crude oil is of two kinds, paraffine base and asphaltum base, and varies from 11 to 46 gravity BaumB. The light-gravity oil does not require heating before use; but oil of less than 20 gravity Baumé should be heated to just below its vaporizing point, in order to reduce its viscosity and to obtain the highest efficiency of combustion. Heavy California oil vaporizes at 230". F., Mexican crude oil at from 175" to 210" F., and Beaumont, Texas, oil at 142" F. Fuel oil vaporizes at 130" F. Scientific Installation Essential The burning of oil is an art based on science, and the "rule of thumb" should never be employed, especially in heating heavy crude oil. Thermometers should be used on both suction and supply pipes so that the fireman or smith will at all times know the temperature of the fuel. The indicators on such thermometers should be large enough to be readily seen by the operator, night or day. It is necessary to have an unvarying pressure of steam or air through the burner, to atomize the fuel. Also, the air used to support combustion should be delivered through a nozzle under the burner at a constant pressure. The oil pressure must not fluctuate; under no eircumstances should it vary more than 1 1b. The oil pump, which should be brass-lined, with the aid of the pulsometer and superheater, will insure a constant

Jan 1, 1916

By William G. [Peryam], Bishop Ottey M., Fred E. Johnson

THE Lepanto Consolidated Mining Company is the largest copper producer in the Philippines, and the only one that makes a concentrate valuable principally for its copper; also, it has the only productive copper mine situated at considerable distance from the sea, and is the only one in the islands operating its own smelter. The Lepanto mine is in the Municipal District of Mankayan, sub-province of Benguet, near the heart of the Mountain Province of Northern Luzon. It is 165 km. from its shipping point at San Fernando, Province of La Union, on the shore of the China Sea. Elevations at the property range from 3200 to 4500 ft. The climate is very mild; temperatures are seldom below 50° or above 80°F. Annual rainfall approaches 170 in. and is mainly confined to a rainy season lasting from May to October. The months from December to April, inclusive, are usually extremely dry. This climatic pecularity entails the problems of landslides on the roads during the rainy season and water shortage during the dry season. The Mountain Trail, an improved rock-surfaced toll road maintained by the Commonwealth Government, gives access to the property. The 105 km. over this road may be traveled in about 4 ½ hr. during the dry season. Very few sections of the road are straight for more than 100 meters, and about one third of it is a one-way road controlled by gates. All freight for the property is delivered by the Manila Railroad Company to Bauang, La Union, 45 km. west of Baguio and at sea level, where it is picked up by freight trucks and hauled via Baguio to Lepanto. Granulated matte is hauled in bulk from Lepanto in four-ton lots on freight trucks to the storage bins at San Fernando. Freight rates from Bauang to the property amount to P8.00 per ton, and from the property to San Fernando the rate is P16.00 per ton. The company pays the road toll, P1.00 per truck, each way. THE CAMP The company maintains a modern camp, well separated from the mine and reduction plant. Comfortable and sanitary living quarters, water, and light, as well as free medical service are provided all employees. A hospital staffed by a doctor and two nurses takes care of all medical treatment. Considering recreational facilities of great importance, the company has provided two tennis courts, two basketball courts, and badminton courts for employees. Bowling alleys and a motion picture theater are operated by the company, the profits being expended for other recreational purposes. A moderate sized recreation hall serves for barrio dances, meetings,

Jan 1, 1942

By L. T. Sanchez, W. T. Rogers

This paper presents the results of a production experiment evaluating the effect of the use of oxygen in the bessemer converter with respect to its relation to blowing time, amount of steel scrap or cold pig iron which can be melted and teeming temperatures of the molten steel. THE simplest and probably the most efficient and economical means for producing steel from molten pig iron is the bessemer converter. This process is essentially a mechanism for bringing oxygen into contact with the molten metal thereby oxidizing the elements silicon, manganese, and carbon. The source of oxygen is that contained in the air and is limited by that amount. In theory, there is an intermediate chemical reaction due to the mass action of iron and air which is represented by the equation: 2Fe + O2- = 2Fe0 The FeO then becomes the oxidizing agent and reacts with the silicon, manganese, and carbon to produce silicon dioxide (SiO,), manganese oxide (MnO), and carbon monoxide (CO), in the order mentioned. The balancing of the various chemical reactions shows that in the ordinary bessemer process more than sufficient oxygen is delivered to satisfy these reactions, which indicates that a further increase in the oxygen content of the air would serve no useful purpose. In spite of this reasoning, however, it will be demonstrated that when melting stelel scrap, commercially pure industrial oxygen introduced into the normal air blast will permit the melting of a larger proportion of steel scrap than is possible with air alone and also will increase the production rate by a substantial reduction in blowing time. It will also be shown that when melting cold pig iron, it is possible to melt up to twice as much pig iron with the use of oxygen as can be melted without added oxygen with a nominal increase in blowing time. At the plant where this study was made, oxygen is purchased from a nearby supplier and piped directly into the plant. It is supplied to the bessemer converter by being introduced directly into the blast line through an automatic control system which may be adjusted by the bessemer operator or "blower." Normally, iron is supplied by five blast furnaces to an open hearth shop containing twelve 165-ton capacity stationary furnaces and a bessemer converting shop having three 30-ton converters. The molten metal from the blast furnaces is allocated to one shop or the other in a manner to best provide for maximum production. At various times, it is desirable to direct the maximum amount of available blast furnace production to the open hearth shop. This has the effect of reducing the production of the bessemer shop. When this condition exists, it is desirable to supplement the reduced iron charge at the bessemer with steel scrap or cold pig iron to the maximum possible extent in order to maintain high total ingot production. This condition becomes aggravated when one of the blast furnaces is taken out of production for any repair job and particularly for extended periods such as when relining is necessary. The purpose of the investigation and experimentation with oxygen was to determine whether or not it would contribute to increased flexibility in the use of the raw materials, molten iron, steel scrap, and cold pig iron. In addition to the effect of oxygen on flexibility of operation, it was also necessary to determine the relation of its use to production rate or "blowing time," final finishing temperature of the steel or "teeming temperature," and steel quality.

Jan 1, 1953

By N. W. Sager, H. W. Mossman

The arrangement and general dimensions of the reverberatory furnaces and waste-heat boilers for the Nos. 2 and 3 smelting units at the McGill plant of the Nevada Consolidated Copper Corporation are shown in Fig. 1. Each of the two reverberatory furnaces has its own set of waste-heat boilers, and before February, 1932, each furnace, when it was in operation, delivered waste-heat gases to its own boilers only. In January, 1933, the idea was suggested of using some of the waste-heat boilers as economizers to operate in conjunction with the waste-heat boilers that were generating steam. The execution of this idea gave rise to the gas-flow scheme depicted in Fig. 1, the advantages of which were obvious, particularly so in view of the fact that for some months prior to the adoption of this scheme the smelting had been confined to No. 2 furnace, with No. 3 furnace and its boilers idle. The waste-heat gases from No. 2 furnace pass through the boilers serving this furnace as steam-generating units, thence through the original exit duct to a point where a bulkhead diverts the gases in a reverse direction through two of the boilers of No. 3 furnace. After the gases have traversed these two boilers, another bulkhead placed in the duct between No. 3 furnace and its boilers causes them to traverse still one more boiler, from which the gas path leads to the stack. In this scheme the last three boilers traversed by the gases, i.e., the boilers behind No. 3 furnace, function as economizers and preheat the feed water for the two steam-generating units directly serving No. 2 furnace. This discussion will give consideration to the results obtained in waste-heat recovery by the scheme shown in Fig. 1, and, further, will compare these results with those obtained before the boilers of No. 3 furnace were set up to act as economizers, but before going into the detail of heat generation, distribution and recovery, will present some additional details of the equipment. All boilers in this waste-heat plant are of the Stirling type. The two behind No. 2 furnace, acting as steam-generating units, of class 13, are 40 tubes wide and rated at 1112 boiler hp., while the three boilers behind No. 3 furnace, acting as economizers, are also of class 13, but 28 tubes

Jan 1, 1934

By H. S. Greismer

Alabama washes a larger percentage of its total coal output than any state in the Union. For producing coking coal, three-compartment jigs are favored; mines providing, steam and commercial coal use single-compartment jigs exclusively. Chemists are as necessary to washing plants as recording gages are to power plants. Experiments in Alabama indicate that the sludge released by coal-washing plants can be deposited in streams and on agricultural lands without disastrous results. CAMPBELL, 1 in 1896 said: "The Birmingham district in Alabama has certain great advantages for there are few places in the world where fuel and ore are so near together, although, unfortunately, both are of inferior quality; the ore being low in iron and high in phosphorus and the coal giving a weak and impure coke." This statement could not have been, questioned in 1894, but by 1896 several Robinson-Ramsay washers had been put into operation and weak and impure coke was a thing of the past. The Warrior coal field of Alabama contains six seams that yield coking coal of excellent quality. In descending- order, they are Brookwood, Milldale, Pratt, America, Mary Lee, and Black Creek. In each of these, however, bands of rock and bone coal of varying thickness are stratified with the coal; and while some of these impurities can be separated from the coal by the miner, much of it must be loaded out with the coal: Consequently, unless the coal is washed before being coked, it will yield a coke of varying ash and sulfur content quite unsatisfactory, for blast-furnace use. At present, the Gulf States Steel Co. is producing run-of-mine coal at its Virginia mine, from the America seam, that is uniformly low enough in ash to permit of coking without washing, but this is the only exception to the above statement. Alabama was the first state to experiment seriously with coal washers, was the first to install washers on a large scale and, up to the present, has maintained the lead. The Alabama State Mine Inspector reports that during 1923 20,919,303 tons of bituminous coal were mined in Alabama; 12,858,499 tons of which were delivered to coal washers for treatment.

Jan 9, 1924

By W. G. Woolf, E. R. Crutcher

The electrolytic zinc plant of the Sullivan Mining Co. is in Government Gulch, at Silver King, Shoshone County, Idaho, in the Coeur d'Alene mining district, about one mile from the Bunker Hill smelter and about three miles from Kellogg. The Sullivan Mining Co., which owns both the zinc plant and the Star mine, is itself owned jointly by the Bunker Hill & Sullivan Mining & Concentrating Co, and the Hecla Mining Company. The Tainton high-density, high-acid process is used at this plant, therefore some parts of its flowsheet differ markedly from those of electrolytic plants using the low-density process. While it is a custom plant and has treated a variety of zinc concentrates originating from various sources in a number of states, it has obtained its main supply of zinc concentrates chiefly from the zinc-bearing ores of the Coeur d'A1ene mining district, of which the principal important sources have been the Bunker Hill and Star mines. Thus the main feed has been zinc concentrates obtained from complex lead-zinc-silver ores. For the year of 1935 the concentrates had the following composition: water, 9.5 per cent; Zn, 50.6; Pb, 3.3; Fe, 9.4; S, 31.0; Cu, 0.2; insoluble, 2.9. Flowsheet The concentrates are delivered to the plant in railroad cars, which discharge into storage bins directly under the track. Bottom-dump cars are dropped directly into the bins while a slip scraper, pulled by an air tugger hoist, is used to empty box cars. There are five storage bins for concentrates, each of approximately 300 tons capacity. At the bottom of each bin there are individual belt feeders that discharge onto a common conveyor belt, which in turn discharges to another belt that carries the concentrate onto a weightometer belt for weighing. Immediately after the weighing, the crushed oversize return from the roasters is added to the load on the belt and then the ground-up sweepings, lumps and roaster cleanings. An electromagnet at the end of this conveyor removes tramp iron from the feed. The belt discharges to a set of 30 by 12-in. rolls, which break up lumps or frozen chunks of concentrate, etc. Beneath the rolls there is a screw conveyor, which conveys the roll discharge to a

Jan 1, 1937

The following letter has been received from the Chairman of the United States Shipping Board. August 1, 1918. American Institute of Mining Engineers, New York, N. Y. GENTLEMEN: I am going to call upon your organization for some teamwork. The time has come for Americans everywhere to put themselves solidly behind American ships. Our railroads must no longer stop at the ocean. We are building an American merchant fleet of twenty-five million tons-three thousand ships. We are backing modern ships with modern port facilities, establishing our bunkering stations all over the globe and will operate with American railroad efficiency. We will carry American cargoes at rates corresponding to our railroad rates-the cheapest in the world. Fast American passenger-and-cargo liners will run regularly to every port in Latin America, the Orient, Africa, Australia. Are you taking steps to use these ships to increase your own prosperity? Do you realize that American products of factory, farm and mine can be delivered to customers in foreign countries on terms which will build lasting trade? Do you realize the possibilities for bringing hack raw materials to extend your products and trade? We must all take off our coats and work to bring these American ships home to the people of every American interest and community. The manufacturer must think of customers in Latin America as being as accessible as those in the next state. The farmer must visualize ships carrying his wheat, cotton, breeding animals, dairy products and fruit to new world markets. The American boy must think of ships and foreign countries when he chooses a calling.

Jan 9, 1918

By C. W. Burrows

The burden a man can endure depends on its magnitude and the number of times it is applied, as well as on many other factors. The resisting power of steel likewise is dependent on many factors. The maximum load it can support indefinitely is much less than its possible momentary burden. A light load straining the metal far within its elastic limit, if repeated indefinitely, ruptures the metal. The magnitude of the stresses that may be alternately applied and removed without rupturing the material depends on the number of repetitions of the cycle; the intervening periods of rest; the thermal, electric, and magnetic conditions, as well as the presence of secondary streessses. While we are unable to formulate the law between the magnitude of repeated stresses necessary to produce rupture and the length of time during which these alternating stresses may be applied before rupture, my experience is that the time increases as the magnitude of the force decreases, according to a law which! to a first approximation, is exponential. In other words, F = Ce-t where F= magnitude of alternating stresses; t= time for which stresses are continued; C = constant, which depends on the conditions under which the operations are carried out. It is quite probable that the same type of law holds for the magnitude of a continuous force and its duration. When a rod of steel is subjected to the blows of a drilling hammer several things happen. First, one end is struck by the hammer and the other end is driven into the rock, thus causing a bodily displacement of the entire material of the steel drill. The greatest stresses involved in this bodily displacement are located at the two ends of the steel and manifest themselves by the distortion of these parts. Second, after the blow has been delivered, the steel drill, due partly to its own resiliency and partly to the resiliency of the rock, experiences a vibratory motion as a whole, this motion is quite similar to the bouncing of a rubber ball thrown violently upon a smooth pavement. Third, the free vibration of the drill is determined only by its own natural period of vibration. This motion involves all the characteristics of any vibrating body emitting sound. Certain elements of the rod will experience alternations of condensation and rarefaction and, in these regions, the individual steel

Jan 1, 1922

By Benj. F. Tillson

To most mine operators, it seems evident that there is a drill-steel problem, although under certain conditions the amount of drill-steel breakage does not appear serious. What is at fault? It may be one or a combination of circumstances. The development of rock drills to the hammer-drill type in place of the old reciprocating piston drill, probably is one important cause for the greater steel breakage. Perhaps the manufacturers of drill steel have failed to realize what alloys are needed for this new service or have overlooked the changes in the art of rock drilling. Although suitable alloys are provided, they may not be SO handled in the manufacturing process as to be in the best condition to withstand the demands of rock drilling. The mine operator may be at fault in desiring to use the smallest possible drill-steel section in order that the gage of the drill bits may be correspondingly small and the amount of footage drilled per unit of labor may be greater; or the black-smithshop practice at the mines may need improvement. The miner who uses the drill steel may also require more intensive supervision and education. Again, the manufacturer of rock drills may have failed to study the types of blows the steel alloys will withstand satisfactorily and which their tools are delivering and so he may not know whether or not a different design of rock drill would equal or excel their present drilling speeds without treating the drill steel so severely. All of these hypotheses probably have their supporters. On the other hand, perhaps we are seeking too much service from drill steel and we need a fuller realization of the fatigue strains developed and should prepare to relieve these by a periodic heat treatment of the steels. However, we do not find any proof that any of these suppositions may be held responsible for the drill-steel breakage attendant to mining. The proposed investigation of this matter by the U. S. Bureau of Mines is fully warranted in the promotion of the conservation of labor and material, and the safety of the workmen. The great range of field conditions, as well as the scope of the research, requires the energetic cooperative support of many interests. Various manufacturers of steel,

Jan 1, 1922

By L. O. Bacon

IN the spring of 1952 Calumet and Hecla Inc. began a geophysical program near Shullsburg, Wis., in the Wisconsin-Illinois lead-zinc district, to assist the geological and drilling exploration programs. Sphalerite, the ore mineral in this area, occurs with a considerable amount of marcasite. In general the higher the percentage of sphalerite in the ore, the higher the percentage of marcasite; however, there may be mineralized zones high in marcasite but low in sphalerite. Orebodies in general are relatively narrow and elongated. Reconnaissance exploration of virgin areas consisted of drilling along lines at 500 to 1000-ft intervals. Whenever a mineralized zone was intersected additional drilling was done to determine whether or not the zone carried zinc percentages of ore grade and, if so, to delineate the body. Calumet and Hecla Inc. was interested particularly in a method that would show the direction and position in which to drill from the original discovery hole. This suggested a method of contacting the orebody, which occurs at depths ranging from 100 to 200 ft, with one electrode. After a number of procedures had been considered it was decided to use equipotential line methods. Instrumentation and Field Procedure: The power source used was a 400-cycle Signal Corps motor-generator having a rated output of 1200 w at 115 v. The detection unit consisted of test probes, a two-stage amplifier, and &apos;head phones for determining the equipotential line by audio null. Field procedure consisted of contacting the sulfide body in the hole by a specially built spring-loaded expanding electrode which would be lowered into the hole, in the closed. position, until it reached the mineralized zone. At that point the mechanism was released so that the arms would make contact with the walls of the hole in the same manner that a well-bore caliper does. This electrode was then connected to the generator, the other terminal of which was connected to the surface electrode. The surface electrode was either a point, a straight line, or a circular line electrode. The point surface electrode was placed at a considerable distance from the hole, that is, 2500 to 5000 ft, depending somewhat upon the depth to the sulfide zone and the distance from the hole containing the subsurface electrode to which exploration was considered. A higher voltage output from the power source would have been desirable, although no difficulty was encountered in determining null points to a distance of 500 ft. Straight line electrodes used were approximately 1000 ft in length at a distance of approximately 1000 ft from the hole containing the subsurface electrode. The circular line electrode used consisted of 4400 ft of bare No. 10 copper wire, which provided a circle of 700-ft radius. All line electrodes were fastened to the ground at 20-ft intervals with copper staples 4 in. long. Equipotential points were determined at 25 to 100-ft intervals. Location of the points was made by plane table and stadia. In open country with a three-man crew, the points were plotted as they were determined. With a two-man crew, pegs were placed at equal potential points and surveyed in after several equipotential lines were determined. Although interpretation of results given in this paper is qualitative, the paper as a whole would not be complete without a discussion of the theoretical distribution of potential due to a circular line electrode. The theoretical distribution would be required if a quantitative interpretation of the measured potential distribution were to be made. The theoretical potential distribution due to a circular line electrode can be obtained by means of Legendre polynomials. Using the method as developed by Weber,&apos; the potential for a circular line electrode in a semi-infinite, homogeneous, isotropic medium can be expressed, using the spherical coordinate system, as: V(r,n) = pi/2pR[S(r/R)nPn(cos?)Pn(cosa)]r<R [1] er V(r,a) =pI/2pr[S(r/R)Pn(cos?) Pn(cosa)]r>R [2] where: 0 is the angle at the origin, taken on the axis of the circular electrode between the axis and the line drawn to the circumference of the circular electrode. a is the angle between the axis and the radius, r.

Jan 1, 1957

By A. J. de Witte

It has been observed that upon invasion of a sand containing oil or gas and connate water by mud filtrate the hydrocarbons are more rapidly flushed by the filtrate than is the connate water.&apos; In time, it appears that the following displacement pattern emerges: oil (or gas) is being swept ahead by connate water which, in turn, is pushed by invading fluid. Eventually the connate water may have "banked up" sufficiently to form a zone of appreciable thickness leading the invading front. In the extreme case, Fig. 1 shows how the situation will develop.&apos; Immediately adjacent to the borehole (radius r) there is a zone, the invaded zone proper, where the connate water has been flushed out completely and which contains only residual oil or gas and mud filtrate. This zone (extending out to a radius ri) grades more or less abruptly into the next (thickness A), which contains only residual hydrocarbon and connate water. Beyond it, again more or less sharply bounded, extends the virgin formation with the original interstitial water saturation Sw. The various saturations are indicated in Fig. 1. The three zones generally will be marked by resistivity contrasts owing to their different fluid contents. As the connate water is usually more saline than the invading fluid, the second zone having a high connate water saturation forms a concentric cylindrical ring or annulus of low resistivity R, around the borehole.&apos;,&apos; It will be referred to as the "low zone." Fig. 1 shows schematically the resistivity profile. It is clear that the phenomenon of a low zone could not occur in invaded water sands. The presence of a low zone, therefore, would be a qualitative indication of a hydrocarbon-bearing formation. Granted that the phenomenon is real, if it were pronounced enough to be detected, one might thereby have a means of locating oil or gas in the ground. This is the aim of the "displacement logging" method.&apos; In any case, the presence of a low resistivity zone will affect the reading of electric logs. Using conventional log interpretation techniques, one must be aware of this and, if necessary, correct for it. Whether correction or detection should be the goal will depend primarily on the magnitude of the effect. The type of log which is likely to be affected most is the induction log. The current pattern of an induction log is concentric with the borehole. Any concentric ring of low resistivity, therefore, will tend to become crowded by current lines. This zone of maximum current density will obscure the relative contributions to the current conduction by other portions of a bed making, for instance, the contribution of the virgin formation less significant and thus detracting from the value or the induction log as an R, reading device. The present note is intended to ascertain the maximum possible effect of a low zone in the extreme case portrayed by Fig. 1 on three commercially available types of induction logs, sc., 5 FF 27, 5 FF 40 and 3 F 60. Type 5 FF 27 or 5 PF 40 is currently run in conjunction with a 16 in. normal and SP log. This combination is gradually replacing the regular electric survey (two normals and one lateral) as a standard log in many areas. On the basis of the scheme of Fig. 1 (sharp boundaries between successive zones), the thickness of the low zone may be computed for various amounts of infiltration measured by the ratio of invaded zone diameter to hole diameter D,/d. Referring to Fig. 1, we can set up a material balance for the water displaced from the invaded zone proper and the water present in the low zone. The volume of connate water displaced from invaded zone upon flushing by mud filtrate is: V1 = e p (r2 - r2) ? Sw . where c = bed thickness ? = porosity Su, = original connate water satu-ration. The volume of water finally accu-mulated in the low zone is: Vz = e p [(r1 -?)2 - r2 4] ? Sw . where SW1 = 1 — Sor = "water" saturation in the invaded and low zones both. The volume of water originally present in the low zone is:

Jan 1, 1958

By R. V. Ramani, T. V. Falkie, T. E. Wilson

In order to increase the efficiency of underground coal operations, it is necessary to understand not only the operating system in detail, but the interrelationships between system components and the operating environment. With such an understanding comes more exact application of simulation and other operations-research techniques to plan and operate mining systems. One aspect of the coal-winning system in need of better definition, namely, the performance of continuous miners, is dealt with in this paper. Detailed discussion of this research will be found in a report by the authors. Mining Rate The performance of continuous miners-that is, the rate at which miners can deliver coal from an underground face to the outby transportation system-is dependent upon the geometric and geological conditions of mining, the characteristics of the mining machinery, and the skill of the machine operators. In order to predict mining performance, it is necessary first to identify these factors, secondly to devise a data-collection technique, and lastly, to select a method of data analysis for drawing meaningful conclusions. Factors Affecting Mining Rate: The factors that affect the mining rate can be conveniently grouped in the following four categories: Geological Factors-The natural conditions such as seam thickness, hardness, and cleavage. Geometrical Factors-The height and width of the entries and the angle of drivage with respect to the main entries. Machine Factors-The mechanical characteristics and physical dimensions of the machine, condition of the machine, condition of the bits, and related items. Other Factors-Operator efficiency, size consist of cutter product, etc. A classification such as given here is not rigid; some of the factors can be included in more than one of the categories and some can be combined as ratios. Data Collection The empirical nature of this investigation required measurements of variables under operating conditions in underground coal mines. A majority of the variables measured were continuous over a specific range of values; for example, the "width ratio" varied in practice from 0.25 to 1.00. "Dummy" values were assigned to certain operational variables in order that the variables would have numerical expression, i.e.: [ ] Data Analysis The method of analysis was to test a large model containing potentially significant variables and then reduce the model to one containing only the essential significant variables. The procedure of "scanning" a large proposed model and subsequently condensing the model consisted of two cycles of the same series of computational procedures. Correlation analysis, factor analysis, and regression analysis were the three statistical procedures that were used to derive the rate-generating equation. Correlation and Regression Analysis: Multiple linear correlation coefficients between all of the variables were calculated using the Pearson product moment formula along with the statistical significance of the correlation

Jan 1, 1975

By Frank Firmstone

. In the early decades of the past century the method of starting iron blast-furnaces by " scatlolding-down" seems to have been in universal use for coke-furnaces and, at least in this country, for charcoal-furnaces also. It was likewise used at some of the early anthracite-furnaces, as is stated in the important " Reminiscences" printed by the late Samuel Thomas.&apos; It had been generally superseded in Great Britain as early as 1850, by some variations of the plan now used, and was rarely used when Percy wrote.2 On the Continent it was still followed ill the 60&apos;s; and was elaborately described by Valerius in 1852. I have never read any statement as to when and by whom " scaffolding " was first dispensed with, but the following extracts from journals of my father seem to fix an authentic date for an instance in this country: " [Pittsburg] July 12, 1838.—Went with Major Wade to his foundry,etc., and he gave me a letter to D. Tyler, Esq., Farrandsville, Agent Boston Coke & Iron Co." " [Lockhaven, Pa.] July 18, 1838.—Heard there was a flat-boat going to take some gentlemen up to Farrandsville. Got on boat and to Farrandsville, 5 or 6 miles, in 1 hr. 30 min. Dined, and then delivered my letter to Ma Tyler, who showed me much attention, and everything belonging to the furnace, and had the water turned on the wheel that I might see the blowing-cylinders at work. There are two cylinders, 6-ft. stroke, -56 in. in diameter, but they find that one makes more than enough blast for the furnace. The furnace has 16-ft. boshes, 7-ft. tunnel-head, and is 47 ft. high, 3 tuyeres, and 3 heating-ovens [hot blast-ovens]. A

Jan 1, 1908

By R. V. Smith

The theory for unsteady radial flow of gas, as simplified by Aronofsky and Jenkins, has been reviewed and compared with actual well performance. This study indicated that the theory should be modified by the introduction of an empirical "rate of flow" function "Y(q)". The expansion of the theory on the flow of gas to include a Y(q) term bridges the gap between the theory of unsteady-state flow of gas and actual gas-well behavior. Apparently, the Y(q) term is a function only of the rate of flow for a given well. The completion factor or skin effect is associated with the Y(q) function in such a manner that at least two or more sets of drawdown or pressure build-up tests are needed to separate the Y(q) function and the completion factor. Since the Y(q) function used in this report represents an energy loss dependent on rate of flow near the well-bore of a gas well which is in addition to the loss required by Darcy&apos;s law, the Y(q) function is related to the exponent of the back-pressure curve for a gas well. Techniques and equations are presented which permit the estimation of stabilized and "short-time" deliver-abilities. INTRODUCTION The published theory of radial unsteady-state flow of gas through porous media may be divided into papers1-" which use Darcy&apos;s law as a premise and other papers4-7,16-10 which recognize that Darcy&apos;s law may not hold for flow into gas wells. Miller8 compared theory based on Darcy&apos;s law with actual gas-well behavior and concluded that flow into gas wells deviates from Darcy&apos;s law. Houpeurt4,5 in a thorough investigation of flow into gas wells concluded that the flow rate into a gas well is proportional to the difference between the squares of upstream and downstream pressures elevated to a power between 1.0 and 0.5. Houpeurt also concluded that the deviation from Darcy&apos;s law is caused by irreversible, kinetic energy exchange between the flowing gas and the porous media. Tek9 presented a complete evaluation of the flow problem in gas wells. Several investigators state that the inclusion of gas compressibility into flow equations based on Darcy&apos;s law does not begin to explain the flow problems suggested by gas-well behavior. For many years, natural-gas engineers have realized that gas wells by actual test exhibit behavior which cannot be explained entirely by Darcy&apos;s law. Many tests on gas wells have indicated that a value of 0.85 for the exponent in the back-pressure equation is a fair average for many wells. However, Darcy&apos;s law either for steady or unsteady-state conditions indicates that the exponent should be 1.0. The purpose of this paper has been to examine this deviation from Darcy&apos;s law and to present a reasonable mathematical means of evaluating the deviation. The method presented in this paper combines previous theory with an empirical function to provide a description of gas-well behavior. All equations in this report are for constant terminal-rate-of-flow conditions. While present methods of testing wells usually are not based on this concept, they are believed to approach the constant terminal-rate conditions closely in many tests. However, several gas wells were tested by holding the rates of flow constant in order to approach more closely the constant terminal rate of flow; the results confirm the concepts presented in this paper. REVIEW OF THEORY FOR UNSTEADY-STATE FLOW OF GAS The basic partial differential equation for the flow of natural gas has been given in technical literature; the equation for a cylindrical reservoir with the well in the center is

By R. M. Hardgrove

PULVERIZED coal was first used for firing cement kilns about 45 years ago, with such success that it has continued in general use. Based on cost, pulverized coal is usually the most economical fuel. In some localities natural gas or oil is cheaper, but 85.5 per cent of all cement kilns at present are fired with pulverized coal. The pulverized-coal flame is well suited to kiln conditions. Its soft, luminous nature gives it a high radiating property and the length of the flame can be readily varied to suit conditions in a kiln of any length. Ash and slag cause most of the difficulties incident to the burning of coal. In a cement kiln, however, the ash, being similar in composition to the raw mix and being deposited uniformly in the mix and clinker, occasions no difficulty as long as it remains constant in quality and quantity. Until 6 years ago almost all cement kilns using pulverized coal employed the bin system. The coal was dried and pulverized in a separate plant and delivered into a bin at the kiln. Thus the bin system requires driers, storage-type pulverizers, conveying systems, individual kiln bins, pulverized-coal feeders and primary air fans. The direct-fired system requires only a pulverizer and a blower. Fig. I shows typical plants. The first installations of direct-fired pulverizers were of the impact type but fineness of pulverization was both inadequate and variable. About 1927 a few direct-fired installations employing tube-type pulverizers were made, which were much more satisfactory. As more efficient pulverizers were developed from 1934 on, equipment for direct firing was installed at a large number of plants, until now 196 of about 900 kilns in use in the United States are direct-fired. Most of these replaced antiquated bin systems. PROCESS OF BURNING CEMENT CLINKERS The process of burning cement clinkers is fundamentally one of heat transfer. In boilers and other heat-exchange apparatus the area of the surface receiving heat from the gases is the best unit for comparing output. This seems to be equally true for kilns. The inside kiln surface is the heat-receiving surface and its area is used as the unit of kiln size in the following comparisons. Heat input is based on British thermal units in the fuel. E. G. Bailey,1 in his work on boiler-furnace heat absorption, has used the unit "B.t.u. available." This is the practical calorific power of the fuel, plus the heat in the air used for combustion. In other words, it is the heat available for useful work in the furnace or kiln. The unit of kiln output is barrels per hour. The composition of the raw mix, the percentage of water in the slurry of wet plants and hardness of burning all affect the output. These conditions, however, are hard to evaluate and probably have less effect on the output than irregularities of raw feed and clinker discharge and inexpert control of combustion. WET VS. DRY PLANTS The study presented here covers 88 plants, of which 41 are wet and 47 dry. Of

Jan 1, 1941

By C. M. Loeb, T. E. Norman

THE use of ball, rod and tube mills for grinding ore, cement and other materials has grown so rapidly during the past forty years that the world&apos;s annual consumption of ferrous grinding media for these mills is now estimated to be between one half million and one million tons per year. Ferrous grinding balls constitute the major portion of this tonnage. Obviously they represent sufficient value to justify thorough studies&apos; of the factor sgoverning their performance. The selection of grinding balls is governed principally by: I. Quality (wear resistance, impact resistance, soundness, and the like). 2. Sources of supply and delivered cost. 3. Grinding characteristics or efficiency in the ball mill. This paper deals principally with the quality of ferrous grinding balls. In the study of these factors certain data relative to the fundamental nature of ball wear in ball mills have been obtained. These data are also presented and discussed briefly. THE DEVELOPMENT OF A SUITABLE WEAR TEST A study of the fundamental factors governing the quality of grinding balls has been hampered seriously by the fact that a competent test has, in the past, involved the purchase of several hundred tons of balls of a specific type which were then run in one or more ball mills for a period ranging from several months to several years&apos; duration. Often, during the period of test, it became necessary to change operating conditions or the character of the ore fed, to the test mill with the result that the rate of ball consumption changed and the test figures became of little value. Under such circumstances, progress in the development of better grinding balls, has been necessarily slow. Economic factors and variations in the quality of balls produced by different sources of supply generally make it necessary for each mill operator to determine for himself the most suitable type of balls for his mills. In our own ball mill grinding operations at Climax, Colo., we were faced with this problem. After we had run a few large scale wear tests at considerable expense we decided to investigate the possibilities of a small scale wear test which would be capable of testing numerous types of balls within a relatively short time. The most important requirement of any test is, of course, that it give results which can be used to predict accurately the wear in full scale operations. It was known that Ellis and his associates1,2 had developed a method of testing grinding balls by small scale tests run at the Ontario Research Foundation. Ellis&apos; method of testing was used as a starting point in our investigations. In the course of our tests a number of modifications of the original method were found to be desirable so that, by an evolutionary procedure, a method of small scale

Jan 1, 1947

By Stanley D. Michaelson

THE selection of the proper size of grinding mill for a given installation has long been a subject of discussion by mill manufacturers, consulting engineers, and their clients. It would be presumptuous for the author to propose a simple cure-all. Instead this paper describes a method that has been used over a period of years to determine mill size with a reasonable degree of accuracy. Furthermore, this method claims no theoretical background, for ball milling, as practiced today, is still an art. The selection of the proper size and type of mill must, therefore, be made on an empirical basis. A ball mill, in order to grind, must deliver power through its grinding media to the material to be ground. Assuming that a mill is properly supplied with grinding media, ore pulp, and power, the amount of grinding done is closely proportional to the power expended. With this as a basic premise we can, for the moment, forget about type of mill (i.e., overflow, low-level mill, or other design) and base our calculation of mill size on the total power required to grind a ton of a particular ore from a given size of feed to a given size of product. When this is determined the type and size of mill, its speed, and the weight of the load of grinding media can be selected to absorb this amount of power. DEVELOPMENT OF TECHNIQUE Some years ago, W. L. Maxson, F. Cadena, and F. C. Bond, working in the Minerals Research Laboratory of Allis-Chalmers Co., Milwaukee, developed and described a laboratory technique for comparing grindability (or relative ease of grinding) of various ores at different mesh sizes of grind.1 Additional papers2 have described the work done since that time, and the most recent paper3 summarizes results for hundreds of tests on many ores and materials. The essentials of the testing method are: a I a by 12-in. laboratory mill operating at a constant speed of rotation, using a constant volume of ore, and a constant ball load having a relatively constant grinding surface. Tests are run at 250 per cent circulating load with the feed, circulating load and products screened on calibrated Tyler standard screens. Certain ores from operating mills, representing various mineralogical types, were selected for reference and designated "standard ores." In most cases, detailed studies of the grinding circuits were made when samples of mill feed were taken for grindability tests. Each analysis formed a basis for comparison of the grindability in the laboratory with actual power requirements for a particular grinding circuit. In addition, a critical study was made, to evaluate the differences in the various grinding circuits so that two or more field installations having different layouts (single-stage grinding, two-stage grinding, or others), and using mills of various diameters operating at various speeds with differing circulating loads, etc., could be compared. Large-scale tests by mill operators complemented by laboratory investiga-

Jan 1, 1945

By H. S. Fleming

My original intention was to give a full account of Southern furnaces, ores and cokes; but, owing to the difficulty of getting reliable information, I confine this paper to a general description of the ores used in, and immediately around, Chattanooga. Mr. L. S. Colyar, Manager of the Rising Fawn furnace, wrote me in answer to an inquiry as to what ores they were using, " Below I give you analyses of the mixture of ore we have been using for the past seven or eight years. We haven&apos;t a complete analysis of any ores, but I will give them to you as I have them. Attalla Soft Rising Fawn Hard Rising Fawn Red Ore. Red Ore. Soft Ore. Iron,.... 50.60 38.48 46.82 Silica,. .. 11.69 7.32 22.69 Phosphorus,. .. 1.06 .319 .350 This is a fair average of all the ores we get in this section." The following analyses of Birmingham soft ore were received from the Morris Mining Co., of that place. No. 1. NO. 2. No. 3. No. 4. Iron.....51.60 54.88 51.04 51.20 Silica,. .. 19.32 10.60 17.32 17.08 Alumina,....19 .88 1.18 Phosphorus,. ..129 .287 .116 .464 Mine No. 1, second vein, shows: iron, 46.40; alumina, 4.92; phosphorus,.212; silica, 19.60. Mine No. 1, by analyses obtained from the Mary Pratt Furnace CO., shows: Iron. Silica. Fine red ore,.49.00 18.38 Lump red ore,...54.42 17.10 Bottom vein,.50.00 21.38 Bottom vein,.41.24 13.52 A lot of this ore delivered in Chattanooga gave : iron, 53.32; alumina, 2.10; phosphorus,.529; silica, 18.03.

Jan 1, 1887