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By N. A. MacLeod

Ball wear in wet grinding mills has been the subject of considerable discussion in the last few years. Its importance to millmen is obvious in view of the fact that ball wear may cost from 2 to 4 per ton of ore for the more fortunate operators and many times this cost when a refractory ore must be reduced to a fine state of comminution. Consequently, any factual information that may be obtained on the nature of ball wear may be applied to introduce changes leading to lower ball consumption and therefore also to lower costs. In 1940, while employed with the Britannia Mining and Smelting Co., Britannia Beach, B.C., I had occasion to investigate factors that might affect ball consumption. We performed a number of tests, which will be briefly described. These tests originated as a result of changing from a 3½-in. ball to a 2-in. ball of different type of manufacture and were motivated by the desire to obtain the best possible type of 2-in. ball. Our testing was helped by the fact that the company operated a one-ton electric furnace, so that it was quite easy to make steel changes. It should be emphasized at the beginning that these tests produced results that in their general application to milling should be interpreted qualitatively rather than quantitatively. As will be evident in the description of the method used in testing, the balls were subjected to an intermittent batch grinding treatment, which may have produced results in rate of steel wear that may not have an accurate correlation with steel consumption in full scale mill operation. However, with this bed and rather arbitrary method of testing, results were obtained that may be of some interest and may stimulate further work on this subject. Unfortunately, owing to pressure of war work, our tests, which we had intended to repeat on a continuous grinding operation, had to be discontinued. Method of Testing Tests Were made in a Small cylindrical mill, rotating at 72 r.P.m. Internal dimensions of the mill were 10½-in. dia. by 9 in. The cylindrical inner surface was lined with rubber and the two ends were formed by steel plates. A charge of 2000 grams dry feed and 2000 C.C. of water was added evev 2 hr., the mill being drained of pulp before the addition of each charge, by tilting and sluicing with water. The mill was run On afternoon and night shift for daily periods of about 15 hr. total grinding time, and each test lasted from 60 to 150 hr. This batch method of testing involved a higher average pulp temperature in the mill, which undoubtedly affected comparative results in Some of the tests. Hardness oF Ball vs. Wear In these tests, special "eutectoid" steel balls were used, some of which were chill-Cut and quenched, some sand-cast and quenched, and some sand-cast and annealed to a dead soft condition. The results obtained, using as mill feed an ore containing considerable free quartz and

Jan 1, 1947

By H. R. Stahl

The ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classi- fiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9½-in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 52¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1¾ in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges

Jan 1, 1947

By H. R. Stahl

The ore of the Southeast Missouri lead district consists essentially of galena in dolomitic limestone. The galena usually is in a very finely disseminated condition. The experience of recent years in the concentrating plants of the St. Joseph Lead Co. has shown that dry crushing carried to a point where about 90 per cent of the ore passes a 14-mesh screen liberates the larger portion of the galena, and minimizes further necessary reduction in rod or ball mills. Concentration practice has attempted to carry size reduction to a point where economic unlocking of the galena is obtained, without causing undue sliming of the very friable galena. While 60 per cent of the tonnage of the district is treated by the flotation process, only 40 per cent of the total concentrates is produced by that means, as efforts have always been made to keep the grade and tonnage of flotation feed as low as possible. The milling practice of the district for some years has recognized the generally accepted idea of better selective grinding in rod mills as compared with ball mills, with the attendant decrease in production of fines. The function of rod and ball mills in the various concentrating plants is the grinding of table middlings and coarse tailings that carry lead values too high to be discarded. The simplified flowsheet in Fig. I illustrates the place of grinding mills in the circuit. The arrangement of classi- fiers, tables and mills constitutes a closed circuit within somewhat flexible limits. In 1937 experiments were conducted in replacing the ball charge in a 6 by 4-ft. ball mill with rods 2 in. in diameter and 51 in. long. This step was taken with some hesitancy, as it was easy to picture a mass of tangled and crossed rods. However, no difficulties developed, and after a short trial period, six more ball mills were converted to rod mills. These mills are of the spur gear and pinion, belt-driven type, using in most cases a 75-hp. motor and having a 9½-in. open trunnion discharge. As ball mills, they carried a charge of 16,000 to 20,000 lb. of balls, 2 in. in original size. The feed to the mills was dewatered by dewatering wheels and introduced at 70 to 75 per cent solids, with a single scoop feeder. One change in mill design was found advisable when rods were introduced. The existing end liners had an outward flare of 3½ in. in 12 in. These were replaced by linings having vertical interior faces, in order to prevent undue end travel of the rods. This change gave an effective working length of 52¾ in. inside the mill. No alterations were made in the shell liners, which are of the smooth type, made of manganese steel 2 in. in original thickness. The rods at present in use are 51 in. long and 1¾ in. in diameter. The composition has been varied at times; at present they contain 0.70 per cent carbon and 0.70 per cent manganese. Some breakage and etching occurs as the rods become small, but this is considered preferable to using softer rods, which may kink and cause tangling of the charge. The mill charges

Jan 1, 1947

By F. Cadena, Fred C. Bond, Walter L. Maxon

Grinding is one of the major problems in present-day milling practice, and in many cases, it represents one of the main items of expense. It becomes necessary from time to time, to compare grinding practice in one plant with that in another plant, and in making such comparisons, it would be very desirable were there available accepted standards of grinding resistance for the materials involved. Such a set of grinding constants would also be of considerable value in planning new grinding installations. There have been some tests published, notably those of Lennox,' which have been generally cited as a basis for such comparisons. Most of the tests made have been based upon open-circuit grinding, in which a weighed sample of the ore was crushed to pass a certain mesh, and was then ground under certain conditions, either wet or dry. Grinding was continued until a certain desired fineness was attained, and the comparisons based on the grinding time, or the samples were all ground for a certain specified length of time and comparison based upon the screen analysis of the products. Such methods have been acceptable where the ore was to be ground in open circuit. However, in modern milling practice, many grinding problems involve closed circuits wherein the oversize returns to the mill and a large circulating load builds up. A method of testing that does not take the circulating load into consideration might yield results quite different from those to be expected in practice, because the minerals in the ore that offer maximum resistance to grinding will accumulate in the circulating load up to an equilibrium value. Thus the nature of the ore charge in the mill at any instant may differ greatly from that of the new feed, with a consequent error in the rate of grinding determined by such a method. For example, let it be assumed that an ore to be tested consists of equal parts of difficult grinding mineral and easy grinding mineral. This is to be compared with an ore used as a standard, and consisting of minerals which are all of about the same grindability. If a comparison

Jan 1, 1935

By E. W. Edson, F. W. Hertal

In some fields the problem of mud fluid is simple and easy of solution. But in the Ventura Avenue field the acquisition and disposal of good drilling mud is not the least of the problems that confront the operator. In fact, mud fluid in the Ventura Avenue field is a very important factor in the drilling of the wells. It is quite probable that many of the factors affecting mud in this field will not be found in most of the fields with which the reader is familiar. However, in the fields where the mud fluid causes no apparent problem it is probable that some of the points brought out in this paper could be used to economic advantage. In fields where caving and squeezing formations and high gas pressures are present, as in the Ventura Avenue field, it is believed that application of methods outlined will be of great assistance in drilling wells. On account of the sandy nature of the formation and the lack of suitable shale bodies in the strata drilled through in the Ventura Avenue field, the wells make practically no mud during drilling. It is necessary therefore to supply drilling mud to the wells that will carry the drill cuttings to the surface, and also will " wall up" the drill hole and prevent the sandy formations from caving. Mixing and Handling Mud Fluid The clay used to make the mud fluid is obtained from a shale stratum in the upper Pico (Plioccne) formation found about a mile on either side of the field. Since high gas pressures and caving formations were first encountered in the Ventura Avenue field and mud fluid was needed to combat them, the mixing of this mud fluid has gone through a process of evolution. First, the clay was mined and hauled to the derrick, where it was shoveled into the well and left for the churning action of the bit to mix into a mud in the hole. The next move was the construction of a mixer to mix the clay to a mud fluid which was pumped into the well. These mixing

Jan 1, 1930

By Richard S. McCaffery, Joseph F. Oesterle, Oscar O. Fritsche

The method of obtaining the data on which this paper is based is explained in detail in the paper on Determination of Viscosity of Iron Blast-furnace Slags.' The work was made possible by a grant of funds from Engineering Foundation, and it gives the results now obtained, continuing the general subject of our previous work and reported in former papers2 The data experimentally obtained are in the form of a temperature-viscosity curve for any slag of particular composition. (See Fig. 1.) From these curves, the viscosity of a slag of a definite composition and at a definite temperature is taken off and is plotted on a plane triaxial diagram showing compositions of silica lime and alumina on a plane representing constant-magnesia content (Fig. 4). The temperatures at which these results were plotted are 1300°, 1350°, 1400°, 1450°, 1500°, 1550" and 1600" C. and since the constant magnesia plane sections were an even 1 per cent. magnesia apart, magnesia analyses not of even unit magnesia composition were projected to these even unit magnesia planes. For example, 1.78 per cent. magnesia and 2.35 per cent. magnesia were projected to the 2 per cent. magnesia plane. The triaxial diagrams on which this first plotting was made were drawn on celluloid sheets, a transparent material on which it was easy to draw. When all the experimental work was plotted on celluloid in this way, the sheets representing a definite temperature were stacked in the proper order and were examined by transmitted light. Where

Jan 1, 1932

By Will H. Coghill, J. Bruce Clemer

Flotation has been so closely allied with the sulfide minerals and their early and associated oily reagents that the term "oil flotation" has erroneously been applied to the entire flotation process. Today, the term "chemical flotation" might be used to advantage. Flotation literature has dealt primarily with the removal of sulfides from non-sulfide gangues, and the separation of one sulfide from another. Now, the application of the process to the separation of nonsulfide minerals is coming to the fore; and from the standpoint of the number of commercial minerals involved, the flotation of nonsulfides offers the larger field. Here, soap or soap-forming substances, such as the fatty acids and their derivatives, are used; hence there may well be a subdivision of the chemical flotation process called "soap flotation." Early experimenters learned that it was easy to form a froth or foam that would carry a part or all of the ore charge over the lip of the flotation machine; but this could not be regarded as flotation because of the lack of selectivity. Doubtless soap was one of the first reagents tried, and doubtless it was rejected on account of its gangue-carrying properties. Now this same gangue-carrying ability makes soap an acceptable reagent, for we are dealing with that which, among sulfides, is considered "gan-gue." As in the flotation of sulfides, so in the flotation of nonsulfides, the selective flocculation of the mineral and the dispersion of the gangue are the first objectives. In fact, these two phenomena epitomize flotation. They depend upon the unlike surface properties of the respective minerals; and as long as the components of an ore have unlike surface properties, a separation is possible, at least in theory. Often in nonsulfide flotation it makes little difference which mineral floats and which is depressed; generally only a separation is sought.

Jan 1, 1935

By J. A. Holmes

Venezuela has continued during 1929 to demonstrate that it is destined for many years to come to be one of our most important sources of crude petroleum. The producing fields of the Bolivar Coastal district along the east side of Lake Maracaibo have continued their phenomenal production, and two new fields have reached the near producing stage, while the results of deep test wells have strengthened the predictions and expectations that oil will be found and produced from many other pools and from several different geological formations. The finding of these new oil pools is not going to be easy, because the surface evidences of geological structure are scanty, and the subsurface problems are difficult, but intelligent effort backed by ample resources may be counted upon to overcome these obstacles and lead to a continued increase in the number of producing fields in Venezuela. During 1929 the total production of Venezuela was approximately 135,953,000 bbl., an increase of nearly 30,000,000 bbl., or 28 per cent., as compared with the preceding year. During 1929 oil in steel storage in Venezuela decreased from 13,000,000 to about 11,000,000 barrels. Tables 1, 2 and 3 briefly summarize Venezuelan production by fields and by years, and show the trend of production since its establishment on a commercial basis. Table 1.---Production of Crude Oil in Venezuela by Years, 1917 to 1929 All Figures in Barrels of 42 Gallons Total PRodUction Total PRoduction Year VeneZUela, Bbl. Year Venezuela, Bbl. 1917....................... 227,000 1927.................... 63,181,000 1918....................... 368,000 1928.................... 106,000,000 1919....................... 258,000 1929.................... 135,953,000* 1920..................... 548,000 Grand Total Since Ineep- 1921...................... 1,544,000 tion................... 380,791,000 1922....................... 2,448,000 Increase, 1929 over 1928.. . 29,953,000 1923..................... 4,170,000 Per cent. Increase......... 28.2 1924....................... 9,234,000 Per cent. 1929 Production 1925....................... 20,219,000 of Total............... 35.7 1926....................... 36,641,000 * Includes estimate for closing period of 1929.

Jan 1, 1930

By J. B. Austin, R. L. Rickett

Measurements of the rate of decomposition of austenite at constant temperature are commonly represented by plotting the percentage transformed on linear coordinates against time on either a linear or a logarithmic scale, a method that yields a curve that in shape resembles an integral sign (Fig. 1). A graph of this kind is not always convenient because extrapolation to the beginning or end of the transformation, or even intebrpolation, may be uncertain unless there are a large number of observations, which seldom happens; furthermore, the degree of consistency of a set of measurements is difficult to judge from such a curve. It is desirable therefore to find some mode of plotting, or even better to find a standard variety of graph paper, that gives a straight line instead of the integral shape. The straight line has many advantages; the number of readings required to fix the position of the curve is greatly reduced, interpolatiorr is easy and rapid, extrapolation can be carried out with greater confidence and, finally, the consistency of the measurements is readily estimated from the scatter of the points. The last advantage is particularly valuable when the data are obtained by direct microscopic examination alone, because they are influenced by the skill of the observer in estimating the relative size of areas, possibly also by undetected inhomogeneity of the specimen. A search for ways of representing the observations as a straight line yielded two methods which are described in this paper. The paper also includes an account of some interesting relations brought to light in applying these methods to the available experimental data. Usually the best way to derive such a method of plotting is to base it on established principles of chemistry and physics, as has been done, for example, for the temperature variation of the equilibrium constant of a chemical reaction. But so little is known of the mechanism of the decomposition of austenite, and so little is available in the way of well established background for changes of this kind, that it is simpler to rely upon purely empirical means. The general plan therefore is to study existing data with a view to finding some function of the direct observations that will

Jan 1, 1939

By J. B. Austin, R. L. Rickett

Measurements of the rate of decomposition of austenite at constant temperature are commonly represented by plotting the percentage transformed on linear coordinates against time on either a linear or a logarithmic scale, a method that yields a curve that in shape resembles an integral sign (Fig. 1). A graph of this kind is not always convenient because extrapolation to the beginning or end of the transformation, or even intebrpolation, may be uncertain unless there are a large number of observations, which seldom happens; furthermore, the degree of consistency of a set of measurements is difficult to judge from such a curve. It is desirable therefore to find some mode of plotting, or even better to find a standard variety of graph paper, that gives a straight line instead of the integral shape. The straight line has many advantages; the number of readings required to fix the position of the curve is greatly reduced, interpolatiorr is easy and rapid, extrapolation can be carried out with greater confidence and, finally, the consistency of the measurements is readily estimated from the scatter of the points. The last advantage is particularly valuable when the data are obtained by direct microscopic examination alone, because they are influenced by the skill of the observer in estimating the relative size of areas, possibly also by undetected inhomogeneity of the specimen. A search for ways of representing the observations as a straight line yielded two methods which are described in this paper. The paper also includes an account of some interesting relations brought to light in applying these methods to the available experimental data. Usually the best way to derive such a method of plotting is to base it on established principles of chemistry and physics, as has been done, for example, for the temperature variation of the equilibrium constant of a chemical reaction. But so little is known of the mechanism of the decomposition of austenite, and so little is available in the way of well established background for changes of this kind, that it is simpler to rely upon purely empirical means. The general plan therefore is to study existing data with a view to finding some function of the direct observations that will

Jan 1, 1939

By Leo P. Tarasov

Magnetic torque studies of ferromagnetic single crystals have been carried out in a number of laboratories during the last decade&apos;s2 and some work has been reported on polycrystalline material showing preferred orientation. Goss3 used the method extensively in developing electrical sheet with desirable magnetic properties, but he used it as a control rather than as a means of studying the texture. Bozorth4 made a quantitative correlation of a torque curve he obtained from some cold-rolled iron with the texture as determined from X-rays, but did not extend his studies beyond the one specimen. Other studies, which have been reported from time to time2,5 have also been incomplete in the sense that no effort was made to study a whole series of samples, differing only in degree of cold reduction or in the annealing temperature, in order to ascertain the effect of such treatments on the texture as revealed by torque measurements. It is the purpose of this investigation to make such a study, mainly in order to see how useful the method can be made when used under favorable circumstances and to find what kind of a correlation can be expected between magnetic and X-ray studies of texture. Before proceeding further, it is desirable to say a little about single crystals, so as to make clear the concepts involved in magnetic torque measurements. Magnetic Torque Method It is fairly well known nowadays that single crystals of ferromagnetic substances have directions of easy magnetization and of hard magnetization. In Fig. 1 are reproduced typical magnetization curves for a single crystal of iron, which show that in a < 100 > direction saturation occurs at low fields compared to those required to saturate the same crystal in a <110> or a <111> direction. (The magnetization intensity I is defined by the equation B = H + 4pI.) At fields high enough to

Jan 1, 1939

By Leo P. Tarasov

Magnetic torque studies of ferromagnetic single crystals have been carried out in a number of laboratories during the last decade&apos;s2 and some work has been reported on polycrystalline material showing preferred orientation. Goss3 used the method extensively in developing electrical sheet with desirable magnetic properties, but he used it as a control rather than as a means of studying the texture. Bozorth4 made a quantitative correlation of a torque curve he obtained from some cold-rolled iron with the texture as determined from X-rays, but did not extend his studies beyond the one specimen. Other studies, which have been reported from time to time2,5 have also been incomplete in the sense that no effort was made to study a whole series of samples, differing only in degree of cold reduction or in the annealing temperature, in order to ascertain the effect of such treatments on the texture as revealed by torque measurements. It is the purpose of this investigation to make such a study, mainly in order to see how useful the method can be made when used under favorable circumstances and to find what kind of a correlation can be expected between magnetic and X-ray studies of texture. Before proceeding further, it is desirable to say a little about single crystals, so as to make clear the concepts involved in magnetic torque measurements. Magnetic Torque Method It is fairly well known nowadays that single crystals of ferromagnetic substances have directions of easy magnetization and of hard magnetization. In Fig. 1 are reproduced typical magnetization curves for a single crystal of iron, which show that in a < 100 > direction saturation occurs at low fields compared to those required to saturate the same crystal in a <110> or a <111> direction. (The magnetization intensity I is defined by the equation B = H + 4pI.) At fields high enough to

Jan 1, 1939

By Ott F. Heizer

The mines and the mill of the Nevada-Massachusetts Co. are on the east slope of the Eugene Mountains, in Pershing County, Nevada, 8 miles northwest of Mill City, a station on the Southern Pacific R.R. and on the Victory Highway. Fig. 1. Supplies and, concentrate are transported by truck over an automobile road of easy grade. The property is at an average elevation of 5500 ft. and climatic conditions are generally favorable for continuous operation throughout the year. The ore treated is supplied by two adjoining mines, the Stank and the Humboldt. Physical and mineral characteristics of the ores produced by these two mines are substantially similar, although variations occur in hardness and mineral distribution.&apos; The economic mineral is scheelite (CaWO4), composed, when pure, of 80.6 per cent tungstic trioxide (WO3) and 19.4 per cent lime (CaO).

Jan 1, 1935

By D. W. Philips

The problem of strata control has of necessity been prominent in mining, and accounts of individual experience and research abound in the literature of most mining countries. Organized rescarch began in Britain by the formation of a committee, upon the recommendation of the Safety in Mines Research Board in 1923, to investigate methods of reducing accidents from falls of roof in coal mines. This committee examined current practice in supporting underground workings in the principal coal fields, and its findings have been published in the Board&apos;s publications.&apos; - Toward the end of these investigations it was suggested by the Chief Inspector of Mines that committees of the several mining institutes be formed to examine the problem in their respective districts. The formation of these committees marked a new phase in research on strata control. The work, though being done in the separate coal fields, followed a scheme outlined by the chief Mining Engineer of the Research Board, and was coordinated by his office. FACTORS INVOLVED The subject is being considered in Britain from two aspects: 1. To provide. by supports, means to prevent falls of fractured ground. 2. TO develop methods whereby the strata can be maintained in the least broken condition and thus make falls less likely. The first aspect is the one that will most likely lead to immediate results from the accident standpoint. It depends largely upon the assimilation of already acquired knowledge regarding the properties of supports and their correct application. The second aspect is perhaps more fundamental, and involves also the properties of the strata and their behavior under the influence of the forces induced by mining. The problem involves many natural factors that cannot be controlled, such as depth, inclination and thickness of seam and nature of the rocks. Other factors, which may be considered to a certain extent controllable, cannot be varied readily and with sufficient frequency to facilitate easy investigation; some of these are: rate of advance of the working face, length of face, direction of the face with regard to cleat and inclination, system and arrangements of rock packing behind longwall faces. In any given case the very many factors present have a mutual influence upon the effects produced in the mine. For example, a study of the effects of rate of advance cannot be done without considering the effects of other factors such as nature of strata, direction of face and quality and arrangement of supports. Conversely, the length and direction of a working face

Jan 1, 1947

By D. W. Philips

The problem of strata control has of necessity been prominent in mining, and accounts of individual experience and research abound in the literature of most mining countries. Organized rescarch began in Britain by the formation of a committee, upon the recommendation of the Safety in Mines Research Board in 1923, to investigate methods of reducing accidents from falls of roof in coal mines. This committee examined current practice in supporting underground workings in the principal coal fields, and its findings have been published in the Board&apos;s publications.&apos; - Toward the end of these investigations it was suggested by the Chief Inspector of Mines that committees of the several mining institutes be formed to examine the problem in their respective districts. The formation of these committees marked a new phase in research on strata control. The work, though being done in the separate coal fields, followed a scheme outlined by the chief Mining Engineer of the Research Board, and was coordinated by his office. FACTORS INVOLVED The subject is being considered in Britain from two aspects: 1. To provide. by supports, means to prevent falls of fractured ground. 2. TO develop methods whereby the strata can be maintained in the least broken condition and thus make falls less likely. The first aspect is the one that will most likely lead to immediate results from the accident standpoint. It depends largely upon the assimilation of already acquired knowledge regarding the properties of supports and their correct application. The second aspect is perhaps more fundamental, and involves also the properties of the strata and their behavior under the influence of the forces induced by mining. The problem involves many natural factors that cannot be controlled, such as depth, inclination and thickness of seam and nature of the rocks. Other factors, which may be considered to a certain extent controllable, cannot be varied readily and with sufficient frequency to facilitate easy investigation; some of these are: rate of advance of the working face, length of face, direction of the face with regard to cleat and inclination, system and arrangements of rock packing behind longwall faces. In any given case the very many factors present have a mutual influence upon the effects produced in the mine. For example, a study of the effects of rate of advance cannot be done without considering the effects of other factors such as nature of strata, direction of face and quality and arrangement of supports. Conversely, the length and direction of a working face

Jan 1, 1947

By Henry F. Herley

The New Orient mine, owned and operated by the Chicago, Wilmington & Franklin Coal Co., has caused a great deal of comment and interest because of its unusual features and huge daily production. It is in Franklin County, Illinois, adjacent to the town of West Frankfort. The seam being mined is the Illinois No. 6 which in this locality is 9 to 12 ft. thick and lies approximately 500 ft. below the surface. Without fear of contradiction, it can be said that this is one of the largest coal mines in the world from a daily production standpoint. During 1928, New Orient produced 2,400,000 tons, which amounts to over 10,000 tons each day the mine operated that year. The mine is built for a capacity of 12,000 tons but has actually hoisted over 15,000 tons in one 8-hr, shift. This enormous production can perhaps be better visualized if it is realized that each day&apos;s operation depletes over two acres of coal. The general features of this mine have already been fully covered in a paper by George B. Harrington,&apos; President of the Chicago, Wilmington & Franklin Coal Co. In a paper of the nature of Mr. Harrington&apos;s, it was unnecessary to stress any particular phase of mining conditions, consequently the ventilation problems were touched upon but lightly. In this paper an attempt has been made to describe some of the problems which have been encountered in providing adequate ventilation for the mine. Ventilating Conditions The acreage allotted to New Orient mine has a length, north and south, of approximately 6 miles, and a width, east and west, of 2 to 4 miles. A number of important considerations caused the hoist shaft, air shaft and surface plant to be located near the southeast corner of the property, instead of near the center, as is usual. During the initial phase of the mine development, the mine was ventilated through these shafts, although there was a full realization that at some future date other shafts would be required.

Jan 1, 1930

By David P. Hale

The Cartersville barite district is near Cartersville, Ga., in the southeastern part of Bartow County, about 43 miles northeast of Atlanta. The area over which active mining is being done extends about six miles in a northeasterly direction and is about three miles wide. Barite production in Georgia began about 1907,&apos; but the existence of the mineral was known much earlier. At present four companies are operating washing plants and three other plants are operating on leases from those companies. The chief impurities in the ores are silica and iron oxide. The silica is easily washed and jigged out and gives little trouble. Any gravity separation between the iron and the barite is generally poor, owing to the slight difference in specific gravity. Two operators have turned to magnetic separation as the solution to the problem and have made remarkable improvements in the finished product. By very close sizing prior to jigging, the separation between iron and barite has been greatly improved. In some localities manganese occasionally appears as an impurity, but is fairly easy to eliminate by gravity methods. A considerable percentage of iron and silica is eliminated by being broken free in crushing. Until methods for the recovery of fine ore were introduced a minimum of crushing was done, but modern methods allow more than formerly. The depression years brought about the need for increased recovery of ore handled on the part of the producers and a demand for a higher grade of product on the part of the buyers. Improvements in milling and mining practice have had a far greater effect, however, than merely the production of a better grade of ore with greater recovery. They have opened for mining large areas that a few years ago were considered noncommercial deposits because of their inaccessibility and their high iron content. It is no exaggeration to say that the new methods have more than doubled present known reserves of barite ore of economic value.

Jan 1, 1942

By David P. Hale

The Cartersville barite district is near Cartersville, Ga., in the southeastern part of Bartow County, about 43 miles northeast of Atlanta. The area over which active mining is being done extends about six miles in a northeasterly direction and is about three miles wide. Barite production in Georgia began about 1907,&apos; but the existence of the mineral was known much earlier. At present four companies are operating washing plants and three other plants are operating on leases from those companies. The chief impurities in the ores are silica and iron oxide. The silica is easily washed and jigged out and gives little trouble. Any gravity separation between the iron and the barite is generally poor, owing to the slight difference in specific gravity. Two operators have turned to magnetic separation as the solution to the problem and have made remarkable improvements in the finished product. By very close sizing prior to jigging, the separation between iron and barite has been greatly improved. In some localities manganese occasionally appears as an impurity, but is fairly easy to eliminate by gravity methods. A considerable percentage of iron and silica is eliminated by being broken free in crushing. Until methods for the recovery of fine ore were introduced a minimum of crushing was done, but modern methods allow more than formerly. The depression years brought about the need for increased recovery of ore handled on the part of the producers and a demand for a higher grade of product on the part of the buyers. Improvements in milling and mining practice have had a far greater effect, however, than merely the production of a better grade of ore with greater recovery. They have opened for mining large areas that a few years ago were considered noncommercial deposits because of their inaccessibility and their high iron content. It is no exaggeration to say that the new methods have more than doubled present known reserves of barite ore of economic value.

Jan 1, 1942

By L. P. Davidson

ThE new electrolytic zinc plant of the American Zinc Company of Illinois commenced operation in April 1941. The simple flowsheet using the standard current density and the economic reasons that dictated it are given herewith. The Evans-Wallower Zinc Co. built an electrolytic zinc plant at Monsanto, Ill., in 1929 and operated it for about two years. The plant stood idle from 1932 and was purchased by the American Zinc Co. in 1940. Because of the condition of the equipment, excepting the motor generator set, a complete rebuilding was necessary. The Fairmont City plant of the American Zinc Co. could treat the leach residues, the purification cake and the skimmings from the melting furnace in its various plants. This possibility permitted a simple flowsheet with good economic results. The flowsheet is shown in Fig. I. The calcined concentrates are leached in mechanically agitated tanks with spent electrolyte and such ferric iron solution as necessary. When the acid has been neutralized and the iron coagulated, the fluid is discharged to settling tanks. The underflow from these thickeners is filtered on a Moore filter; the filter cake leaves the plant for retreatment and the filtrate joins the thickener overflow. Zinc dust is added to this solution until the copper and cadmium are precipitated. The pulp is filtered through Shriver presses, which deliver the purified solution to stor- age tanks, and the copper-cadmium-zinc cake is shipped for retreatment and recovery of these three metals. The purified solution is electrolyzed; the regenerated acid is returned for use in the leaching plant and the cathode zinc, after stripping, goes to the melting furnace, from which it is cast in slabs. The skimmings from the melting furnace are shipped to the Fairmont City plant for reduction to metal in the retort furnaces. The choice of this simple flowsheet was made possible only because of the facilities of the Fairmont City plant. The long leaching pried, with high acid content and high temperature which characterize the high-density process, gives a higher zinc extraction but also dissolves more impurities, which must be removed. The same is true to a lesser degree of the multi-step leaching of the western low-density plants. By the sacrifice of some initial extraction, solutions can be obtained that are relatively easy to purify and give good ampere efficiency. This loss in initial extraction is made up in the retreatment of the residues in the Waelz plant. The impurities that accompany the zinc thus extracted are inconsequential when the Waelz product is treated in a distillation plant. It was recognized that this flowsheet would require a somewhat greater amount of sulphuric acid—-eithcr as acid or as sulphate sulphur in the calcine—because of poorer washing. The sulphuric acid plant at Fairmont City could deliver 60" Bè. acid to compensate for the sulphate loss and the cost would be less than the cost of more intensive washing.

Jan 1, 1943

By L. P. Davidson

ThE new electrolytic zinc plant of the American Zinc Company of Illinois commenced operation in April 1941. The simple flowsheet using the standard current density and the economic reasons that dictated it are given herewith. The Evans-Wallower Zinc Co. built an electrolytic zinc plant at Monsanto, Ill., in 1929 and operated it for about two years. The plant stood idle from 1932 and was purchased by the American Zinc Co. in 1940. Because of the condition of the equipment, excepting the motor generator set, a complete rebuilding was necessary. The Fairmont City plant of the American Zinc Co. could treat the leach residues, the purification cake and the skimmings from the melting furnace in its various plants. This possibility permitted a simple flowsheet with good economic results. The flowsheet is shown in Fig. I. The calcined concentrates are leached in mechanically agitated tanks with spent electrolyte and such ferric iron solution as necessary. When the acid has been neutralized and the iron coagulated, the fluid is discharged to settling tanks. The underflow from these thickeners is filtered on a Moore filter; the filter cake leaves the plant for retreatment and the filtrate joins the thickener overflow. Zinc dust is added to this solution until the copper and cadmium are precipitated. The pulp is filtered through Shriver presses, which deliver the purified solution to stor- age tanks, and the copper-cadmium-zinc cake is shipped for retreatment and recovery of these three metals. The purified solution is electrolyzed; the regenerated acid is returned for use in the leaching plant and the cathode zinc, after stripping, goes to the melting furnace, from which it is cast in slabs. The skimmings from the melting furnace are shipped to the Fairmont City plant for reduction to metal in the retort furnaces. The choice of this simple flowsheet was made possible only because of the facilities of the Fairmont City plant. The long leaching pried, with high acid content and high temperature which characterize the high-density process, gives a higher zinc extraction but also dissolves more impurities, which must be removed. The same is true to a lesser degree of the multi-step leaching of the western low-density plants. By the sacrifice of some initial extraction, solutions can be obtained that are relatively easy to purify and give good ampere efficiency. This loss in initial extraction is made up in the retreatment of the residues in the Waelz plant. The impurities that accompany the zinc thus extracted are inconsequential when the Waelz product is treated in a distillation plant. It was recognized that this flowsheet would require a somewhat greater amount of sulphuric acid—-eithcr as acid or as sulphate sulphur in the calcine—because of poorer washing. The sulphuric acid plant at Fairmont City could deliver 60" Bè. acid to compensate for the sulphate loss and the cost would be less than the cost of more intensive washing.

Jan 1, 1943