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By John L. Schroder

INTRODUCTION-SYSTEM DESIGN The design of an underground mining operation requires the integration of transportation, ventilation, ground control, and mining methods to form a system which provides the highest possible degree of safety for mine personnel; the lowest cost per ton of product ; the most suitable quality of final product, considering its intended use; the maximum percent recovery of the reserve, consistent with cost and subsidence rights ; and the minimum disturbance to the environment. The selection of mining method is generally the controlling factor in overall mine design. Some important considerations in the choice of mining method are: 1) Depth of cover over the coal seam. 2) Strength and thickness of rock strata in the overburden. 3) Characteristics of the coal seam, including thickness, nature and location of impurities or irregularities, dip (or pitch), nature and direction of cleavage, hardness, and structural strength of the coal. 4) Nature and strength of the immediate roof (top) over the coal seam. 5) Nature and strength of the floor (bottom) under the coal seam. 6) Presence and amount of explosive gas (methane). 7) Quantity of water likely to be encountered in the mining operation. 8) Any previous or concurrent mining in seams above or below the coal seam to be worked. 9) Any restrictions on recovery required for subsidence control. Topographical and geological reports and maps of the area are usually available from state and federal geological surveys, and contain excellent general information to guide more detailed investigations. Exploratory drilling, and the plotting and analysis of the information thus obtained, are essential steps in any development program, and aid in the correlation and evaluation of information gained by other means. Other good sources of information are operating coal mines that may be in the same

Jan 1, 1981

By L. D. Ricketts

' In recent years the metallurgical field of the copper industry has expanded greatly, the copper ores have become lean add diverse in charactcr, and we are obliged to treat such ores on a very large scale. On the commercial side, the operating and consulting staffs of the great mining companies have been enlarged and organized, and no great capital expenditure is decided upon until it has been investigated and recommended by a highly organized force of specialists as a committee. I believe that in the scientific field and in the most valuable proceedings of our -various societies the old idea of one-man point of view and one-man treatment of principle and detail together has been splendidly developed and should continue, but for the advancement of our art we should deliberately discuss subjects among ourselves in order to do away with the idiosyncrasies of the individual and to gain a composite view reprcscnting the best thoughts of many men. In the development of new ideas leading to the introduction of new methods, the societies could enlarge their present great field of usefulness if such discussion be had, and in my opinion further great good could be done if a number of men of different qualifications and experience would join in preparing a joint book, each dealing with his particular problem, but consulting the other men. In this way individualism would be preserved and various processes would be described, and if properly reviewed the paper would be palpably capable of analysis, and the student would have at his command the associated efforts of a group of qualified engineers. I have neither the time nor the qualifications to make an investigation and study in detail of hydro-metallurgical processes, but for some years I have foreseen the development of such processes, and I have a deep interest in the subject and have given some of its phases careful consideration from a long-distance viewpoint. I have assisted in directing many experiments and in reviewing the merits of many processes, and am glad to make some introductory remarks with the understanding that other gentlemen present will freely and candidly enter into the discussion and help to fill out the few bare bones that I can uncover.

Jan 1, 1916

By Charles R. Wraith

The property of the Roan Antelope Copper Mines, Ltd., is situated at Luanshya, north central part of Northern Rhodesia, and is connected with the main line of the Rhodesia Railways, Ltd., by a 24-mile branch line from N'dola. Most of the material and supplies that are not available in Africa are shipped into the country via Beira, on the East Coast of Africa, a distance of 1473 miles from Luanshya and the nearest ocean port. The output of copper is shipped by rail to Beira and thence by boat to the market. Passenger travel is mostly via Cape Town, a distance of 2162 miles from Luanshya. Excellent passenger trains operate on regular schedules between Cape Town and Northern Rhodesia. Luanshya is 13" south of the equator at an elevation of 4000 ft. above sea level, and has a mild healthful climate. During the rainy season, from October to March, the rainfall averages 53 in. The area is profusely covered with vegetation. An ample supply of water for industrial and domestic purposes is available throughout the year. The more important supplies for smelting purposcs, such as wood, coal, limestone, silica, quartzite, and clay for "dolly" mud, are conveniently available. Excellent grades of coal, coke and firebrick are obtained from the Wankie Colliery Co., Ltd., Southern Rhodesia. The Wankie colliery has just completed the erection of equipment for the manufacture of silica brick. An ample supply of quartzite of good grade is available for this purpose. Silica brick is also obtained from the Belgian Congo and from India. Limestone is quarried from a large deposit a few miles from Bwana Mkubwa, which is 30 miles by rail from the smelter. An abundant supply of clay for " dolly " mud and for lining launders is available within a short distance of the smelter. Metallurgy Copper is contained in the ore as chalcocite, chalcopyrite and bornite. Chalcocite predominates in the eastern end of the orebody almost to the exclusion of other copper minerals. The amount of chalcocite decreases with a corresponding increase of chalcopyrite and bornite toward the western extension, where the latter predominate.

Jan 1, 1934

By H. M. Chance

RECENT research work has shown that coal can be produced, at reasonable cost, from almost all coal-mining districts containing not more than 3 to 8 per cent. of ash. From coal so produced, an abundant supply of coke, suitable for blast-furnace operation, can be made containing from 12 down to 6 per cent. or less of ash. These statements apply broadly to the Appalachian coal belt, including most of the mining districts in Pennsylvania, Ohio, West Virginia, Kentucky, Tennessee, and Alabama. The exceptions will generally be found among those coals carrying a prohibitive percentage of sulfur. Any gravimetric heavy fluid method of separation will effect a reduction in the percentage of sulfur, but this reduction may not be sufficient to produce, from coal unfit for metallurgic use, a product that can be so used. When a large part of the sulfur is in the form known as "organic sulfur, "or is present as very .finely disseminated and homogeneously distributed pyrite, possibly it cannot be eliminated from the low-ash coal produced. When the sulfur exists as pyrite in distinct layers, bands, or nodules, or when it is finely disseminated but is irregularly distributed, often a low-ash coal low enough in sulfur to make good metallurgic coke can be produced from relatively impure coals. It has generally been assumed that the quality of coke on which blast furnaces must be operated depends on the geographic location off the furnace, and that the grade of coke on which a blast furnace must operate is that produced from the best coal obtainable by careful mining; or if the coal is prepared by washing, hand picking, or other means by which the foreign impurities are removed, that the grade of coke on which the furnace must operate depends on what has been termed the "inherent ash" of the particular coal.

Jan 6, 1924

By B. A. Eaton, K. E. Brown, C. R. Knowles, D. E. Andrews, I. H. Silberberg

This paper presents the resitlts of an investigation of two-phase, gm-liquid flow in horizontal pipelines. Experimental data were taken in three field-size, horizontal pipelines, two of which were constructed for this purpose. The data were obtained using water, distillate and crude oil separately as the liquid phase, and natural gas as the second phase. Experimental pressure-length traverse, liquid holdup and flow-pattern data were collected for each set of flow rates. These data were used to develop three correlations that are presented herein. The liquid-holdup values correlated with various flow parameters without regard to the existing flow pattern. The same was true for the energy-loss factors. A new flow-pattern map is presented that appears to be quite reliable, but not required for the pressure-loss calculations. The liquid-holdup correlation and the energy-loss factor correlation are used in conjunction with a two-phase flow power balance, developed during this study, to predict the pressure losses that occur during gas-liquid flow in horizontal pipelines. A recommended calculational procedure is given, as well as a statistical analysis of the results. This procedure lends itself to computer application, since several small pressure decrements are needed to calculate a pressure-length traverse. The correlations are shown graphically, but may be curve fitted with existing curve-fitring computer programs. INTRODUCTION Due to the frequent occurrence of gas-liquid flow in pipelines and the desire to accurately calculate the pressure losses that occur in these lines, two-phase flow is of considerable interest to the petroleum, chemical and nuclear industries. In the petroleum industry, gas-liquid mixtures have been transported over relatively long distances in a common line due to the advent of centralized gathering and separation systems. Long two-phase flowlines are usually accompanied by large pressure drops which influence the design of the system. Gas-lift installations are designed on the basis of known tubing pressures at the wellheads. The horizontal flowline connecting the wellhead and the separator system must be correctly sized in order to minimize the horizontal flowline pressure losses and the wellhead tubing pressure. Practically all oilwell production design involves horizontal two-phase flow in pipeknes. All of the flow processes of oil and gas production must be studied simultaneously to insure good well design. Since the beginning of offshore oilfield development, long horizontal flowlines have been constructed. Because pressure losses greatly influence the performance of producing wells, a method is desired that can be used to predict such pressure losses and select optimum flowline size. Several types of gas-liquid flow exist, and many of these are discussed by Gouse The study of pressure gradients, fluid distributions and flow patterns that occur in horizontaI multiphase flow is made difficult by the great number of variables involved. The various flow regimes give rise to changing velocities of the fluid particles in all directions. These instabilities of the interface between the gas and liquid prohibit the determination of actual vector velocities of fluid particles in each phase. Also, it is practically impossible to arrive at correct sets of boundary conditions. Therefore, most investigators have concluded that a solution to the problem by the classical fluid dynamics approach, whereby the Navier-Stokes equationsM are formulated and solved, is far too complex. Other methods must be utilized to develop general correlations that will predict the behavior of gas-liquid horizontal flow systems. Multiphase flow studies have sought to develop a technique with which the pressure drop can be calculated. Pressure losses in two-phase, gas-liquid flow are quite different from those encountered in single-phase flow; in most cases an interface exists and the gas slips past the liquid. The interface may be smooth or have varying degrees of roughness, depending on the flow pattern. Therefore, a transfer of energy from the gaseous phase to the liquid phase may take place while energy is lost from the system through the wetting phase at the pipe wall. Such an energy transfer may be either in the form of heat exchange or of acceleration. Since each phase must flow through a smaller area than if it flowed alone, amazingly high pressure losses occur when compared to single-phase flow. Most investigators of horizontal two-phase flow phenomena have chosen to separate their experimental data into several groups of observed flow patterns or regimes.

By Donald E. Cummings

MEASURES appropriate for the control of the silicosis hazard in mining cannot be formulated precisely, but sufficient has accumulated during the past quarter century to permit the suggestion of useful, though partly empirical, rules for its alleviation. The industrial hygienist today recognizes that the peril incurred by the inhalation of harmful dust is a function of two variable factors-the degree of harmful exposure and the specific susceptibility of the exposed individual to pulmonary injury. Control of the danger arising out of prolonged contact with pernicious air- borne particulate matter is consequently best exercised by influencing both of these recognized variables. Obviously the problem has two broad aspects-one, medical; the other within the province of the engineer. I-FACTORS COMPRISING THE DEGREE OF INDIVIDUAL SUSCEPTIBILITY Thorough investigations conducted among miners who have labored for years in the presence of recognized silicosis hazards usually reveal an impressive difference in the injury incurred by men whose exposure to dust has been nearly identical. After a quarter century underground, one individual is short of breath and incapacitated with miners' complaint, while a fellow workman suffers no pulmonary embarrassment and is in good health. Physicians ascribe this oft-observed difference in reaction to individual susceptibility and express a universal regret when they deplore the fate that chose to make the predisposed youth a miner rather than a farmer. To trust the health of miners to fate at the present time is even more unfortunate, because many of the factors that influence man's susceptibility to dust are known, and can be accurately determined by competent medical methods. Among the most important factors comprising susceptibility to injury by dust are the anatomical peculiarities of the individual's respiratory mechanism, his physiological response to the inhalation of dust, his previous dust exposure, the existence of pulmonary infections or other pathological processes in his lungs, his age,

Jan 1, 1938

By M. D. Cooper

For the purpose of obtaining some first-hand data in regard to the shooting down of coal in bituminous mines, it was the writer's good fortune to be employed as a shot-firer for almost one year. In all, 6020 shots were fired by him during this period. All of the work was done in what is known as the South Main section of the No. 2 Mine of the Ellsworth Collieries Co., Ellsworth, Pa., a subsidiary of the Lackawanna Steel Go. Position of Shot-firing in the Cycle of Operations The established practice of the Ellsworth Collieries Co. calls first for the undercutting of the coal by electric chain machines across the face of the entry or room. This is followed by replacing any posts removed by the machine-runners, or the setting of new posts or timber, as may be required. After this has been done the machine cuttings are loaded into cars. Then the coal is spragged, and a shot hole drilled to the depth of the undercut. Next in order, the shot-firer loads the hole and fires the charge. The working place is then cleaned up, timbered if necessary, the sides and face squared up, and loose material taken down. The face is again ready to be undercut. Materials Used Permissible explosives were used exclusively in blasting coal. These were Carbonite No. 2, a nitroglyccrine explosive; and three different makes of ammonium nitrate explosives, Tunnelite, Red H, and Mine-ite 5-D. The last-named was used during the greater part of the period. It is interesting to note that even at the beginning of this work, the gases resulting from the explosions of the charges caused no especial distress on. the shot-firer's part and this continued to be the case except on the occasions when a change was made from one class of explosive to another.

Jan 1, 1918

By George S. Rice

The A.I.M.E. Ground Movement and Subsidence Committee, proposed in 1920, held its first technical meeting in February 1923, under the able chairmanship of Mr. H. G. Moulton. The following list of papers and discussions, indicating the scope of the problems and relation to various conditions and method of mining, were given at that meeting by the distinguished mining engineers: J. Parke Channing, on Subsidence at Miami, Arizona; Howard N. Eavenson, on Mining an Upper Bituminous Seam after a Lower Seam Has Been Extracted; J. J. Rutledge, on Subsidence in Two Oklahoma Coal Mines; Benjamin F. Tillson, on Caving Cracks at the Franklin Mine, New Jersey; Dr. F. W. McNair, on the Dome Theory; William Kelly, on disadvantageous effects in Mining and Back-filling Working Upward a Steep-Pitching Iron-ore Bed; Eli T. Conner, on the good results of Back-Filling by Hydraulic Means in Preventing Destructive Subsidence in the Anthracite District; Thomas H. Claggett, on the merit of Systematic Pillar Drawing in Preventing Cracks in an Upper Seam; Henry H. Otto, successful results in the anthracite district in Mining Lower Seams before Upper Seams; Douglas Bunting, on Leading Factors Incident to Successful Mining of an Overlying Seam; Edwin Ludlow, on there being "less damage to buildings when there was complete removal of coal with complete packing than when room and pillar is used without extracting the pillars"; Edward O'Toole on "bumps" in eastern Kentucky mines, which he thought were the result of geologic stresses: H. A. Buehler, on Upheaval of Surface in an Illinois Case in Advance of the Face; and R. Dawson Hall, on The Mechanics of Subsidence from Mining. The author of this paper discussed various kinds of problems of ground movement and subsidence in both metal mining and coal mining; the production of great landslides by workings near an outcrop, as at Turtle Mountain, Alberta, Canada; and general

Jan 1, 1940

By George S. Rice

The A.I.M.E. Ground Movement and Subsidence Committee, proposed in 1920, held its first technical meeting in February 1923, under the able chairmanship of Mr. H. G. Moulton. The following list of papers and discussions, indicating the scope of the problems and relation to various conditions and method of mining, were given at that meeting by the distinguished mining engineers: J. Parke Channing, on Subsidence at Miami, Arizona; Howard N. Eavenson, on Mining an Upper Bituminous Seam after a Lower Seam Has Been Extracted; J. J. Rutledge, on Subsidence in Two Oklahoma Coal Mines; Benjamin F. Tillson, on Caving Cracks at the Franklin Mine, New Jersey; Dr. F. W. McNair, on the Dome Theory; William Kelly, on disadvantageous effects in Mining and Back-filling Working Upward a Steep-Pitching Iron-ore Bed; Eli T. Conner, on the good results of Back-Filling by Hydraulic Means in Preventing Destructive Subsidence in the Anthracite District; Thomas H. Claggett, on the merit of Systematic Pillar Drawing in Preventing Cracks in an Upper Seam; Henry H. Otto, successful results in the anthracite district in Mining Lower Seams before Upper Seams; Douglas Bunting, on Leading Factors Incident to Successful Mining of an Overlying Seam; Edwin Ludlow, on there being "less damage to buildings when there was complete removal of coal with complete packing than when room and pillar is used without extracting the pillars"; Edward O'Toole on "bumps" in eastern Kentucky mines, which he thought were the result of geologic stresses: H. A. Buehler, on Upheaval of Surface in an Illinois Case in Advance of the Face; and R. Dawson Hall, on The Mechanics of Subsidence from Mining. The author of this paper discussed various kinds of problems of ground movement and subsidence in both metal mining and coal mining; the production of great landslides by workings near an outcrop, as at Turtle Mountain, Alberta, Canada; and general

Jan 1, 1940

By J. A. Ede

The property of the Spring Valley Coal Company, situated in Bureau county, Ill., comprises something more than 30,000 acres of coal-lands, on which have been opened four mines, designated as Nos. 1, 2, 3 and 4. Figs. 1, 2, 3 and 4 are plans of the first three; and Fig. 7 is a view of the surface-plant at No. 4, the workings of which it is not necessary to illustrate, as this mine has been practically merged in No. 2. No. 4 shaft is only 500 feet from No. 2, and both are ventilated by the same fan, and are in the same circle of air-supply. In speaking of No. 2, we now include No. 4. The system pursued in all the mines is " long-wall advancing," and the shafts are all vertical. The seam worked is locally and commercially known, in the Illinois series, as No. 3; but its position in the geological series is that of No. 2. It averages in thickness about 42 inches, arid has no partings, but mines in rectangular lumps, which stand a large amount of handling before breaking. An analysis made from Colliery No. 1 gave the following : Per cent. Moisture,.......... 1.45 Volatile combustible matter,.47.27 Fixed carbon,.........41.97 Sulphur,.........2.46 Ash,....... .. .. .. 6.85 Plate 6 shows a photographic view of this seam. The staff seen in the view is 37 inches long. The table on page 188, recorded in my office, which I presume to be correct, may be of interest. The bottom is fire-clay, varying in thickness from 6 to 24 inches. It affords what is considered good mining-ground. The roof immediately above the coal is a dark shale, lying in small bands, from 4 to 9 inches in thickness. Between the

Jan 1, 1900

By Arthur C. Green

A group of 13 cities situated in Los Angeles and Orange counties in Southern California is engaged in constructing an aqueduct to carry water from the Colorado River at a point near Parker, Arizona, to a distribution point near Riverside, California. The aqueduct will have a total length of approximately 241 miles, and is designed to deliver 1605 cu. ft. of water per second, or a little more than one billion gallons per day. The aqueduct crosses a barren, desert waste and the project includes approximately 91 miles of 16-ft. tunnel. The large amount of tunnel driving that has been done in this country in the past few years has resulted in some remarkable developments in equipment and methods. The building of the Colorado River Aqueduct,, with its 91 miles of tunnels, has called on the best engineering knowledge in the country. New and better equipment has been designed for all phases of the work, and this equipment has been coordinated and adapted to the various conditions with a high degree of efficiency. The result has been that records for tunnel driving have been set under adverse natural conditions that exceed the estimates for progress under good conditions. All of the tunnels in the main aqueduct, except the Valverde and Bernasconi at the western end, have a finished diameter, inside the concrete lining, of 16 ft. Fig. 4 shows a section of the tunnel with the concrete lining in place. Fig. 5 illustrates the finished tunnel in three typical sections. The standard tunnel grade is 3.432 ft. per mile in the 16-ft. tunnels, but the Valverde and Bernasconi tunnels have a steeper grade and a finished diameter of 15 ft. 3 in. The rough bore of the large main aqueduct tunnels averages 12.95 solid cubic yards of material per foot of advance in timbered sections and 10.76 solid cubic yards per foot in unsupported sections. Tunnel Driving While the Conway mucker has been used on tunnel work for years, a larger and more powerful Conway was designed to meet the require-

Jan 1, 1937

By Nathaniel Arbiter

THE separation of minerals by flotation can be regarded as a rate process, with the extraction of any one mineral determined by its flotation rate, and the grade of concentrate by the relative rates for all the minerals. So regarded, the significant variables for the process are those that control the rates. These variables are of two types, the first describing the ore and its physical and chemical treatment prior to flotation and the second characterizing the separation process in the cells. This paper will examine the variation in rates for a group of separations, will show that a simple rate law appears to govern, and will consider the relation of the control variables to the rates. The use of rate constants for evaluation of performance and efficiency will be discussed.

Jan 9, 1951

By B. J. Kochanowsky

TO improve blasting methods it is necessary to know how the explosive force acts and how rock resists this force. Because of the tremendous power developed within milliseconds and the great number of other factors directly affecting the technical and economic results, an analysis of the fundamentals of blasting theory is difficult. But since the rules used for layout design and for calculations of size of explosive charges are based on theoretical assumptions, complete knowledge of blasting theory has great practical importance in mining. Analysis of Blasting Theory: It is interesting to note the opinion of blasting experts with respect to contemporary blasting theories. F. Stussi; Professor of the University of Zurich, stated: "We do not have enough experience yet to change our army engineering regulations in blasting and base it on new fundamentals. It is our duty to collect more practical data and to do more research in blasting to close this gap." K. H. Fraenkel,2 editor of the Manual on Rock Blasting published in 1953 in Sweden and written by well-known Swedish, German, Swiss, and French blasting and explosive experts, said: "To the best of our knowledge no suitable formulas for civil blasting work are to be found in the American, French or German literature." Present blasting theory is based upon two assumptions. 1) The blasting force of explosive acts in concentrical and spherical form. 2) Rock resistance against the explosive force is directly proportional to the strength characteristics of the rock. The first classical formula based on theoretical fundamental in blasting theory for explosive charge calculation was introduced by Vauban, a military engineer who lived 300 years ago. It was Vauban who proposed the famous formula L = w3 q, where L is the explosive charge, w = line of least resistance, and q = specific explosive consumption proportional to the weight of rock. Later engineers used q as proportional to the strength of the rock. Since Vauban's time different suggestions concerning blasting theory have been proposed. However, the principles stated at that time so affected the thinking of later generations that his formula is still in use and practically unchanged. The first controversy concerned the form of crater. It was found that geological features of rock affected its form. The factor q was analyzed thoroughly by Lares3 and later by Ohnesorge," Weichelt,5 Bendel,6 and others, but the assumption remained that resistance against explosive force is directly proportional to the strength of the rock blasted. The greatest controversy, which has not yet been settled, concerned w. It was noted that w3 is more appropriate for long lines of resistance and w2 for lines of resistance less than 15 ft. Based on the assumption that the explosive force acts concentrically and spherically, spacings between charges were limited to distances not greater than the length of line of least resistance. Sometimes larger spacing is recommended, but this is due to the advantageous geological and physical properties of rock and not to the action of an explosive force as such. In addition to the classical formula, empirical formulas are used widely. These state that the explosive charge is directly proportional to the volume of blasted rock in cubic yards, and the amounts of explosive required are usually expressed in pounds of explosive per cubic yard of rock. Empirical and classical formulas are contradictory. In the empirical formula, but not in the classical formula, explosive charge is taken proportional to all three space axes: line of least resistance, spacing, and bench height. In spite of this contradiction, both formulas give good results. This is possible because as now practiced the explosive charge calculation for heavy burdens need not be highly accurate. Each, open pit or quarry, usually works with a certain relation between bench height and line of least resistance and between charge spacing and line of least resistance. When these relations are changed, however, the specific explosive consumption q changes greatly. This is one of the reasons why the principles on which the formulas are based appear to be incorrect. In addition to the formulas discussed, others exist and are based more or less on the same theoretical

Jan 1, 1956

By Francis Church Lincoln

I. Situation and Surroundings, The Promontorio mine is situated at the northern end of the Sierra San Francisco de Coneto, in the town of Promontorio, Partido of El Oro, State of Durango, Mexico. As shown in the sketch-map, Fig. 1, the nearest railroad station is Chin-acates on the Mexican International Railroad, 82 miles north of the city of Durango. The mine is 16 miles north of the sta-tion by air-line. It is reached by means of a good wagon-road which first crosses the Guatimapé Plain, passing the ranches of San Antonio and San Julian, to Estacion—a distance of about 15.5 miles—and then entering the mountains crosses the Sierra to Promontorio—a further distance of 11.5 miles— making in all a distance of 27 miles by road from railroad to mine. Promontorio is just beyond the summit of the range, at an elevation of about 8,000 ft. above sea-level, or 1,350 ft. above the Chinacates station. The Castillo de San Francisco, the highest peak in the Sierra, has an altitude of 10,000 ft., and the Promontorio road crosses the summit of the ridge at an elevation of 9,000 ft. by a pass just below this peak known as the Puerto del Almagre. The Promontorio mill is situated at Santa Inés, 2 miles by tram or 3 miles by wagon-road to the north of Promontorio and about 600 ft. lower. The Sierra San Francisco de Coneto decreases in height towards the north and west till it comes to an end at the Mel-chor Arroyo on the ranches of Melchor and Ramos. Beyond this arroyo another range, the Sierra de la Candela, begins. In the vicinity of the Promontorio mine the hills are very rugged and the surface is furrowed by ravines which contain running streams during the rainy season-—July, August and

Jan 1, 1908

By R. D. Pehlke, J. V. Gluck

The limiting activity coefficient of zinc in a "binary " solvent has been described for several dilute ternary solutions by combining literature data from recent interaction studies in the terminal regions with thermodynamic data for the central regions. Results are discussed in terms of the Alcock and Richardson random solution model. The observed behavior agrees semiquantitatively with predictions using the model. RECENT efforts in the thermodynamics of dilute ternary liquid metallic solutions have been mainly concerned with the interactions between two dilute solutes, that is, the terminal region at an apex of the composition diagram.'-' A more general case is that of the behavior of a dilute solute over the full range of a ternary system, or, more specifically, the limiting activity of solute i as projected to the k-j binary side of the i-j-k composition triangle. An implication of this is that the interaction between i and j at the k apex should connect in some way to the interaction between i and k at the j apex. A model to characterize various types of expected behavior has been discussed by Alcock and Richardson.9,10 Such behavior can range from a simple linear relation for In yi over the 1-k binary to more complex curved relations which depend on various assumptions regarding the heat of solution in the j-k binary system and the presence or absence of clustering with solute i. Quantitative data to assess these concepts have been limited primarily to the cases of nonmetallic interstitial solutes in ferrous solutions. Alcock, Richardson, and various coworkers9-13 have discussed the behavior of oxygen, sulfur, silicon, or hydrogen in dilute solution over the range of iron + metal systems; however, the only data considered by them for substitutional nonferrous metallic solutions were the results of Elliott and Chipman concerning the activity of cadmium in the Pb + Sn, Pb + Bi, and Pb + Sb systems.9 These data were limited and did not cover the terminal regions in detail. An opportunity to extend the Alcock and Richardson considerations to the dilute solution of zinc in various "binary" solvents has been afforded by combining results of several of the recent investigations of ter-minal-region interactions with previous data from the literature for the same more concentrated solutions. RESULTS AND DISCUSSION The results of the analysis are shown in Figs. 1 to 5 where the log activity coefficient, (In yZn)xzn--o, is plotted vs the composition of the "binary" solvent. In these plots the end points are the values for In yzn in the appropriate binary solution with Bi, Pb, Sn, Cd, or In at the indicated temperature—usually 550°C (823°K). For the terminal regions the slopes were drawn using the first-order interaction parameters taken from various studies reported in the literature. The data points shown in the central regions were obtained by the short extrapolation to infinite dilution of the logarithm of the activity coefficient calculated from thermodynamic studies reported in the literature for the same systems. The sources of the data are indicated on each figure. Second-order effects were not considered. Reasonably smooth curves can be drawn to connect the central region with the terminal regions for the Zn-Pb + Bi system, Fig. 1, the Zn-Pb + Zn system, Fig. 1, and the Zn-Bi + Cd and Zn-Bi + Sn systems, Fig. 2. As shown in Fig. 2, where central region data are available from two independent investigations, the agreement is quite good. The Zn-Pb + Cd system does not appear to be as well behaved as the previous four, Fig. 3; however, this type of behavior has also been

Jan 1, 1970

By John Rogers

The goal of a detailed quantitative understanding of the variables insulfide flotation in terms of established, or new, scientific principles is still distant, although there has been a good beginning in the past 50 years with more rapid progress since 1930. Advances have paralleled closely those in other branches of surface chemistry and the development of a series of experimental techniques such as contact angle, surface area measurement, radioactive tracers, infrared spectra, electron diffraction, monolayer, and electro- chemical methods of investigation. The second editions of books by Sutherland and Warkl ( 1955) and Gaudin2 ( 19571, are detailed accounts of this progress. The complexity of sulfide mineral flotation systems is indicated by Sutherland and Warkl who list 32 variables of flotation, of which only 22 can usually be controlled by the metallurgist. A consequence of this multiplicity of variables, many of which may be specific to each particular ore, is that ways of floating sulfide minerals from new ores are still developed empirically. How- ever, this review will try to show how the chemistry of the function of frothers, collectors, depressants, and activators in sulfide flotation is being steadily unravelled. Chiefly, it will be concerned with the interaction of pure soluble compounds with individual sulfide minerals, the surfaces of which are often not stoichiometric for several reasons discussed by Plaksin.3 As a result, for example, a freshly cleaved galena surface has cathodic and anodic areas which have different degrees of oxidation and different reactivity to- wards collectors like xanthate. Such complicating considerations should be remembered in the subsequent discussion of pure minerals. THE NEED FOR FLOTATION REAGENTS After considerable controversy it has been established with a few exceptions that unoxidized, uncontaminated sulfides will not float. Moreover all flotation plants recovering sulfide minerals use collectors which are removed from the circuit with the sulfides. Gaudin, Miaw, and Spedden4 have discussed the relation between the crystal structure of solids and the native floatability of the exceptions mentioned above. A second question which has been much explored is the nature of the surface on which a collector like xanthate is adsorbed or reacts. Is this the un- changed surface of the sulfide mineral, or does the adsorption depend on the initial oxidation of the sulfide surface ? Sutherland and Warkl and Gaudin2 discuss the work of Taggart5 and others on this question without reaching any other conclusion than that oxygen was important.

Jan 1, 1962

By Dennis K. Mortensen

The Role of the Operator The ultimate objective of any installation is to be able to accomplish the useful work that it was originally designed for. The operator needs a tool that he can both operate effectively and maintain efficiently. The overall design and subsequent installation of a comminution circuit should be a marriage of good engineering practice and practicality. Since the operator must operate and maintain his equipment, his input into the design and installation is vital to its ultimate success. The design of a crushing or grinding circuit is dependent upon many factors including flowsheet design, equipment selection, physical layout, space restrictions, climatic restrictions and considerations for future expansion. The operator must live on a day-to-day basis with the results of the final design and installation. The most satisfactory operation will result when the operator has a major input into each of the criteria that are used to define the final design and installation. The operator referred to is not only the mining company or organization that is having the installation built for its specific use but, more specifically, the person or persons that will actually be responsible for its daily operation and maintenance. Future references to the "operator" will include those individuals with that daily reponsibility. The operator's input into the installation of the circuit is of equal importance to his input into the design. The engineering organization which puts together all of the various design drawings for circuit installation is generally restricted to trying to get all of the details onto a multitude of 24" x 36" two dimensional drawings with a few isometric drawings thrown in. It's almost impossible to get all of the details of a final installation onto all of the drawings, and field modifications are necessary to make everything fit properly. In a multilevel installation, such as a grinding or crushing circuit, it is difficult to envision space requirements to adequately operate and maintain the various pieces of equipment simply by looking at a set o£ drawings. For some designs it is helpful to build a scale model so that a more realistic view of what the installation will look like can be seen. The scale model allows the operator to make whatever changes in design are needed to operate and maintain his equipment most effi-

Jan 1, 1982

By Arnulf Muan, Hobart M. Kraner

Ferromanganese alloys react with aluminu-silica brick in blast furnace hearths and cause the formation of new phases with low refractoriness and consequent failure of the refractory lining. The nature of these reactions is explained on the basis of petrographic observations, thermodynamic data, and Phase equilibrium considerations. Animproued hearth design resulted from these findings. The production of ferromanganese in blast furnaces presents many problems which are not encountered in the production of iron in the same furnaces. One of the most serious of these problems is the rapid failure of refractories caused by their reaction with the ferromanganese metal. The present paper describes such a "hearth attack", its symptoms, causes, and cure. Following an introductory discussion of operating practice in ferromanganese blast furnaces, results are presented of a petrographic examination of samples from a furnace which had to be shut down because of refractory failure. The causes of this failure are then analyzed on the basis of petrographic observations combined with thermodynamic and phase equilibrium data available in the literature. The latter type of data were subsequently used as a guide in recommending improvements in hearth designs which have since proven themselves successful in practical operations. I) FERROMANGANESE BLAST FURNACE PRACTICE It is considered good practice to blow in a furnace on iron and operate it in this way for a period of time preliminary to producing ferrornanganese. During this time, iron generally penetrates the bottom refractories six or more feet below the original working surface of the hearth. Not only does this metal pervade the joints to these depths, but the pores of the refractories are usually also filled by molten iron. The high throughput of the iron furnace provides ample heat to maintain a high temperature in the hearth refractories. This, together with a high ferrostatic pressure, causes porous clay brick of the hearth to shrink. During subsequent use of such a furnace in ferromanganese production, the iron in the pores of hearth brick if; replaced to some extent by ferromanganese. The replacement is usually not complete. This is probably due to the lower metal throughput and lower prevailing temperatures resulting from this in the brick of the hearth bottom. When the ferrornanganese furnace is banked—or when the temperature in the furnace falls due to a furnace delay—-manganese oxidizes even though the atmosphere is largely CO. Considerable expansion accompanies this oxidation of manganese from metal to oxide. The hearth refractories are then subjected to the disintegrating forces of the volume expansion. These forces may in extreme cases be large enough to lift the furnace off its mantle supports as was described in a previous paper.' When cooling takes place slowly, however, and the temperature remains at a fairly high level, the refractories undergo fluxing action by the MnO. This latter reaction and the reduction of SiO2 of the clay brick is the general subject of the present paper. The observations to be described herein were made on a ferromanganese furnace located at the Johnstown plant of Bethlehem Steel Co. The furnace was lined entirely with clay refractories and had operated on iron for 266 days and on ferromanganese for 96 days prior to being taken off because of a hearth wall failure. II) PETROGRAPHIC EXAMINATION AND CHEMICAL ANALYSIS The photograph reproduced in Fig. 1 shws a cross section of the upper part of an 18 by 9 by 4-1/2 in. bottom block removed from the blast furnace. This brick came from a depth of approximately 6 ft below the original working face of the bottom. It will be noted that the 4-1/2 in. dimension is almost the same as when the brick entered the furnace. Three characteristically different zones marked I, II, and III on the print of Fig. 1 are apparent. Results of chemical analyses of the original brick and samples from each of these zones are listed in Table I. Petrographically the various zones were characterized as follows: The central zone marked I was essentially un-reacted brick. The microstructure consisted of fine needles of mullite, and the pores in this entire area were filled with iron. This is characteristic of

Jan 1, 1962

By W. J. Reagan

The problem of increasing output and decreasing percentage of rejections is a vital one in the manufacture of steel of any kind. The making of basic open-hearth steel for use in rolled steel wheels, tires for locomotive and car wheels, and forgings of various kinds, is a highly specialized industry. The increase in weight of railway rolling stock of all kinds, and the high speed that the public demands of transportation facilities, make it imperative that steel for use in this class of material shall be of the highest quality obtainable. During the past few years it has been the writer's good fortune to have opportunity to study the defects in basic open-hearth steel, and also the results obtained by various changes in open-hearth practice, on such defects. The opportunity for study and the details available are rarely found in the manufacture of steel of any kind. It seems reasonable to assume that many of the results determined here can be applied to the production of steel of other types with a decided improvement in quality, a decrease in the percentage of rejections and an increase in output. All of the material described is basic open-hearth steel with the following range of analysis: carbon, 0.50 to 0.85 per cent.; phosphorus and sulfur, 0.04 max.; silicon, 0.15 to 0.35; manganese, 0.50 to 0.75. All of this material is bottom-cast, therefore tapping and pouring temperatures are higher than in top-cast production. Ingots used in this type of work are of standard length with varying diameters. The ingot is known as duodecagonal, or twelve-sided. The standard body length is 88 in. and the hot top length, 14 in. Fig. 1 shows a typical ingot of this type. The diameter varies from 13 to 30 in. Ingots for special forgings, which cannot be made in molds of standard size, are made in molds of various sizes and designs. These ingots are sliced cold, into blocks of various lengths, depending upon the weight of the forging to be made. Fig. 2 shows the ingot after coming from the slicing lathe. Jusf sufficient metal remains between blocks to hold the ingot in one piece while it is being removed from the lathe; then steel wedges are inserted between the blocks and the blocks are broken apart. This affords abundant opportunity for inspection, as the interior of the

Jan 1, 1930

By Walter Crafts, John L. Lamont

An investigation has been made of the multiplying factors for some of the more common alloying and deoxidizing elements for use in calculating hardenability of steel according to Grossmann's method. The factors for manganese, silicon, and aluminum that were reported in an earlier communication2 have been confirmed by microscopic determination of the half-martensite structure rather than on the basis of an empirically determined hardness corresponding to a half-martensite structure. Measurement of ideal critical diameter by microscopic examination was found to be necessary for other alloying elements, and multiplying factors for nickel, chromium, molybdenum, zirconium, vanadium, titanium, and boron have been determined. The nature and order of magnitude of the multiplying factors reflect the effect of the alloying elements on the structural character of the steel. Except for nickel and vanadium, the multiplying factors agree fairly well with previously published results. Good correlation between actual and calculated ideal critical diameters has been found in developing the multiplying factors and in checking the results against fairly simple steels. However, some discrepancies have suggested the need for a more thorough study of low-carbon and complex steels. Testing Procedure The investigation was conducted principally on steels made at the Union Car- bide and Carbon Research Laboratories, Inc., in high-frequency induction furnaces using Armco iron and standard alloying materials. However, a few open-hearth steels were used in establishing the effect of nickel, and the hardenability factor for boron was based primarily on open-hearth steels. All bars were normalized prior to the machining of Jominy hardenability specimens. The Jominy test was carried out under conditions standard for the test.3 Austenitizing temperatures judged to be appropriate for the compositions of the steels were used. Chemical analysis was determined on a sample taken adjacent to the Jominy specimen. Residual amounts of phosphorus, sulphur, nickel, chromium, molybdenum, and copper were determined on only one heat of each group made with the same materials during the same period. "Acid-soluble" aluminum and zirconium rather than total amounts were determined, as it is considered probable that the acid-soluble amount is more representative of the effective alloy. As tin was present to the extent of only 0.002 per cent, its effect was neglected. The method used to determine boron in the steels shown in Table 9* was developed by Galen Porter, Vincent Napoleon, and Thomas R. Cunningham, of the Union Carbide and Carbon Research Laboratories, Inc., and is a modification of J. Naftel's procedure for boron in soils

Jan 1, 1944