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By Victor deWolfe

INTRODUCTION The Henderson Mine, located near Empire, Colorado, utilizes a continuous panel caving system to extract ore as one of the world's major producers of molybdenum. Any mine using a caving-by-gravity technique of mining must rely on closely controlled draw of the caved ore. This control is essential to insure proper caving action, to avoid damaging load concentrations of weight and to minimize the dilution of ore with waste material. Henderson is no exception. Draw control is a major factor in all production planning, from long- range plans to short-range and day-to-day ore scheduling. Draw control is reviewed constantly and administered daily in an effort to optimize production efficiency, ore recovery, and cave management. MINING METHOD The cave at Henderson is massive, moving slowly through large panels that are 244 m (800 ft.) wide by 610 m (2,000 ft.) long. Generally two cave areas are drawn at one time. The areas under active draw vary in size but can be as large as 244 m (800 ft.) by 244 m (800 ft. ) containing 400 draw points. Each draw point contains 45,360 mt (50,000 st) on the average and takes about two and one half years to exhaust. A complete panel is worked for seven to ten years. No pillar exists between panels, but rather a buffer zone of broken ore, or "static face," is left in each panel to be drawn with the adjacent, yet-to-be-caved panel in efforts of minimizing dilution of a working area from an exhausted one. (Figure 1) Production drifts are driven on 24.4 m (80 ft.) centers through the ore body. Between the production drifts are funnel-shaped draw bells on 12.2 m (40 ft.) x 24.4 m (80 ft.) centers to receive ore from the cave. Each bell is accessed by two draw points, one from the production drift on either side, thus forming a 12.2 m (40 ft.) x 12.2 m (40 ft.) draw pattern. Extraction of the ore is by rubber-tired, 3.8 m3 (5 yd3) load-haul-dump equipment. The LHDs then tram the ore a maximum of 49 m (160 ft.) to ore passes. Cave initiation and bell development are done from the undercut drifts which are parallel to and 17 m (55 ft.) directly above the production drifts. Longhole rings are drilled and blasted from the undercut drifts to define the bells and establish the undercut for caving. (Figure 2) DRAW CONTROL Since the cave line at Henderson is constantly advancing, it is necessary to be continually initiating new cave at one end while exhausting it at the opposite end. There must exist, therefore, an angle on the ore-waste contact in the broken rock from initiation to exhaustion. The basic concept of draw control is to keep this angle as smooth and even as possible, particularly at the time of exhaustion. If this is achieved, draw points are exhausted more or less in a line, avoiding pockets of remaining ore surrounded by exhausted areas. These pockets would cause spotty ore extraction at the time of exhaustion, increasing the amount of dilution occurring while introducing the potential for significant weight problems in the production area. To arrive at the desired angle on the ore- waste contact, maximum tonnage percentages are assigned to each row of draw points increasing at 10% or 15% increments (depending on cave size and velocity of draw) working away from the cave line. The available tonnage indicated by these percentages is the maximum allowable tonnage to be extracted from each draw point until the available tonnage percent- age is increased. As the cave moves, these percentages increase for each draw point regularly. However, in general the tonnage drawn from each draw point is kept at about 50% of this allowable maximum in order to maintain adequate available tonnage in the cave to sustain production for seven months if cave initiation were to cease. This available tonnage cushion is a safeguard built into the draw control program at Henderson to accommodate fluctuations in the rate of cave advance. When draw points move past the row of 100% tonnage availability, they are drawn past the desired 50% at the same increments per row until exhausted. (Figure 3) To achieve proper draw control, the number of LHD buckets to be taken from each draw point is assigned daily. The actual buckets taken, which may at times deviate from the

Jan 1, 1981

By J. Philip Neal, Carroll P. Rogers

Feldspars, the most abundant minerals of the igneous rocks, occur in numerous forms and mixtures. The feldspars of commercial significance are found in widely distributed pegmatites as large crystals usually free of iron- bearing impurities and thus suitable for hand-cobbing, and also in larger bodies where the ore contains various types of feldspar inter- mingled with quartz and relatively free of iron- bearing impurities, or at least readily unlocked from impurities present at relatively coarse mesh. The latter type of ore includes some pegmatites, and the ''alaskite" bodies so common in the Spruce Pine district of North Carolina. On the west coast of the U.S. and in Oklahoma there is-r has been--commercial recovery of natural feldspathic sand as a source of alumina in glass and ceramics. "Aplite" in commercial terminology is a light-colored igneous rock with a granitic composition and a fine sugary texture, often banded with readily removable iron impurities. Two areas in Virginia produce aplite for the glass industry: one located in the vicinity of Piney River, Amherst County, and the other near Montpelier in Hanover County. According to Deer et al. (1966) the word feldspar derives from the Swedish feldt or falt (field) and "spath:" chunks of rock appearing in tilled ground overlying granite. Castle and Gillson (1960) indicate a Germanic origin, citing "spat," which is said to refer to any transparent or translucent material which is readily cleavable. The term "spar" has in the past been applied to minerals other than feld- spar, such as barite, calcite, and fluorite. It is correct to call the latter "fluorspar." There is a true barium feldspar, but it is rare and of no present economic importance. What is recognized as feldspar or "spar" by present producers and consumers consists of three silicate minerals which, if pure, would have the formu- las KAISi,O, (microcline or orthoclase), NaAlSi,O, (albite), and CaAl,Si,O, (anorthite). These are almost never found in pure form in nature, but occur together in great abundance in a three-component system. A considerable number of different combined ratios of these three exists, with the exception that isomorphism between potassium spar and calcium spar is very limited. User specifications for feldspar products are largely built around the prevalent naturally occurring ratios. Pas- sages following will refer to "K-spar," "Na- spar," or "Ca-spar." Nomenclature and Descriptive Terms Relating to both the mineralogy and the economics of feldspar are certain terms whose coverage it is useful to specify. Commercially, aplite is a feldspathic rock mined and beneficiated in two locations in Virginia, in which both titanium and feldspar minerals are present-the latter now being the major, if not the only, economic product. Technically, the term aplite when applied to this rock is of questionable accuracy. But, in light of accepted commercial usage, the term will be here used to designate it. Regarding the Virginia rock, Ross (1941) is cited.

Jan 1, 1975

By C. E. Lesher

THIS paper reports a study of the reactivity of 950°F and 1650°F cokes as measured by relative rates of reduction of iron oxides at temperatures up to 2200°F. Previous work cited shows general acceptance of the theory that reduction by carbon is a gaseous reaction, and that kind and character of carbon as well as particle size have measurable effect on the velocity of reaction. As will be shown, the data obtained in this study con firm those conclusions. The work was not designed to examine iron oxide reduction equilibrium, but if reaction velocity be defined as the speed with which "a reaction tends to approach conditions of equilibrium," the data here presented may be considered as a study of reaction rates, and the relative degree of reduction to metallic iron as the measure of reactivity.

Jan 7, 1950

By A. W. Brune

Conventional methods of calculating the duty cycle of a mine hoist are awkward in that a number of items must be remembered or that they must be collected together in order to have all the calculated data in one place. A more convenient way is to tabulate all factors involved, permitting a ready evaluation of the problem. An example illustrates the convenience of such a table for either of two types of drums, cylindrical and conical. Assume the following conditions: The hoisting depth is 600 ft. Balanced hoisting is used. The weights of empty skip and ore are 6000 lbs and 12,000 lbs, respectively. Production is 3600 tons per shift. Net hoisting time per shift is six hours with a rest period of eight sec, top and bottom, per trip. Acceleration and retardation are each six sec. The diameter of the drum is seven ft; its face is four ft. The sheave has the same diameter. The equivalent effective weight of the drum reduced to rope center is 15,400 lbs. The weight of each sheave is 2100 lbs. 1 The rope is 1 1/4-in. improved plow steel weighing 2.50 lb per ft.

Jan 9, 1961

By W. E. Mitchell

ELECTROLYTIC, production of cadmium at the Great Falls plant started in the first part of the year 1925. Prior to that time, an experi¬mental unit had been in operation for a few months during the year . 1922. At present the plant is the largest producer of cadmium, not only in the United States but in the world. Electrolytic cadmium is characterized by its very low content of impurities. The main impurities, which are lead, copper, zinc, and iron, do not aggregate more than 0.05 per cent., leaving a cadmium content better than 99.95 per cent. The consumption of cadmium has increased rapidly since the war, with the growing use in the plating industry. Cadmium plating has come into prominence because of its connection with the automobile and radio industries. As a protective coating, a deposit of, cadmium is superior to zinc. Cadmium is also used extensively in alloys. Important among these is the alloy with copper used in telephone and trolley wire. In proportion of 0.5 to 1.2 per cent., cadmium adds materially to the strength and wearing qualities. Practically all commercial zinc concentrates contain a small amount of cadmium. In the electrolytic zinc process, the cadmium that is dissolved in the regular leaching operation is removed from solution with metallic zinc, generally added in the form of zinc dust. This precipitate also contains the copper taken into solution and the excess of zinc dust, which is required in order to insure a complete removal of all impurities. The precipitate is treated by a separate process to recover the copper, cadmium and excess zinc. In order to make the metals more soluble and to render insoluble such impurities as iron, arsenic and antimony, the precipitate is roasted. The roasting is carried out in gas-fired McDougall furnaces, with a roasting temperature of about 700° C.

Jan 1, 1930

By F. A. Orth

COAL-mine opera- tors face a difficul- ty in notifying men when there is no work. If the miners come to the shaft when there is no work, they must be paid, by agreement, two hours' wages. This involves a considerable loss since it is not uncommon to em- ploy two or three hun- dred men or more at a single shaft. Many mines have solved the problem by using an electric siren. This carries the "no work" message to all of the miners within a ra- dius of from one and one- half to two miles, so they do not trudge uselessly to work and the operator saves the two-hours wage payment. At the Valley Smokeless Coal Co., Johns- town, Pa., such a siren serves as a start and quit- work signal for three mines which previously re¬ quired three individual signal systems. Similar units are used by the Old Ben Coal Corp. mines, Dolly's Smokeless Coal Co., Rhonda Coal Co., Bird Coal Co., and hundreds of others. Many other applications of the electric siren have been found in mines. They are used on underground locomotives, to awaken miners in the morning, to tell them it is time to come to work, to sound the start and quit-work signal, and to warn of fire or accident. The sirens are most generally used to give a start and quit-work signal, fire, and general emergency alarms as well as no-work signals. They require 3 and 5 hp. to operate. The smaller unit has a sound radius of from three-quarters to one mile, while the larger siren can be heard over a distance of from one and one-half to two miles. The 5-hp. unit has been heard for as great a distance as 14 miles, and 10 miles is not at all uncommon. One of the outstanding features of the 3 and 5-hp. units, which is uncommon in the smaller sirens is that they may be adapted to sound code signals. When used to sound a fire or general emergency alarm the siren may tell at the same time where the fire or acci-dent is. When a remote control device is used with a siren it is possible to operate the siren from any number of distant points. Any number of "start" buttons may be placed at various places. By using one "start and stop" button where the remote control is mounted, and one or more "start" buttons at various places about the plant, it is necessary to run only two wires to the various "start" buttons. If "start" and "stop" buttons are used altogether, then three wires must be run.

Jan 11, 1927

By P. Campbell, E. M. Gardy, L. Hachkowski

Introduction The efficient process control of a uranium mill and its associated mining operation requires a nearly continuous knowledge of the uranium concentration in the ore. Generally, the approach is to use laboratory assays of grab samples from the mill feed belt. In some cases, elaborate and expensive systems have been used to ensure random sampling, but even with these systems, mass-balance discrepancies still exist. There is a requirement for an on-line instrument that can give a prompt, accurate analysis of a large portion of the feed stock. The authors have recently evaluated a laboratory system that achieves these goals using neutron activation and delayed neutron counting. The development of the on-line uranium ore analyzer is a consequence of previous work done at the Whiteshell Nuclear Research Establishment (WNRE) (Campbell et al., 1978 and 1981), and is based on the emission of delayed neutrons after the neutron irradiation of fissionable materials. The mechanism of delayed neutron emission has been described (Keepin, 1965), but briefly is as follows. The fission fragments resulting from irradiation are in an excited state. Certain of these delayed fission products, precursors, decay to a more stable state by the emission of a delayed neutron. The delayed neutrons can be divided into six groups with effective half-lives ranging from 0.2-55 sec. It is important to note that more than 50% of the delayed neutrons are emitted within the first 6 sec after irradiation; this has an influence on the design of the analysis instrument. Measurement Cycle The measurement cycle involves a neutron irradiation, a delay time, and a counting period. The laboratory demonstration model of the ore analyzer comprises a californium-252 neutron source, a source chamber, a pneumatic transfer facility, and a neutron irradiation and counting module [(Fig. 1)]. The ore sample is placed in the irradiation chamber. The neutron source is transferred to the center of the chamber for a 5-sec irradiation while the helium-3 thermal neutron detectors are switched off. As the source returns to the storage chamber, the detectors are switched on and the delayed neutrons counted. The total assay cycle is 16 sec. The current instrument design calls for the ore to be cut from the main belt every 16 sec by a mechanical arm and dropped into the analyzer directly, or via another belt. The cycle starts when there is a sufficient mass of ore in the irradiation chamber to depress a switch and stops when the delayed neutron count is finished. The ore is then dropped through the bottom of the analyzer chamber and returned to the main belt. The counts from each cycle are accumulated and the data reduced to give the desired information.

Jan 1, 1985

By Nicholas T. Camicia

Coal is one of our oldest known commodities. Historic evidence indicates that this fossil-formed fuel was used in China and in Wales some 3000 to 4000 pears ago. Coal was discovered in North America in 1679 by French explorers in northern Illinois, and in 1700 a colony of Huguenots came across bituminous coal in Virginia. One might think by this long history of a single commodity that the string was about played out. Quite the contrary, coal is in greater demand than ever before and coal mining today is a modern industry. It is in an unprecedented growth period as the nation-and the world-demands more and more energy coincident with dwindling reserves of other fossil fuels. By the year 2000 the demand for coal is expected to at least double the current level of some 600 million tons annually despite the advent of nuclear power. To meet this challenge, the coal industry is vitally in need of an increasing supply of technically trained and progressive manpower. And, coupled with a serious shortage of engineers and technicians, it represents a rich vein of occupation opportunities-all the way up to top career levels. Today, coal mining is a complex, highly specialized industry concerned with more than just getting coal from the earth and putting it into the furnaces of the consumer. The industry is concerned with the environment, with maintaining a viable ecology. Recognizing the ramifications of global economics, the industry is very much aware of its important position in contributing to a constantly rising standard of living throughout the world. The key to rewarding employment in the coal industry is skill-skill that has been developed by education, training, and experience. None of these alone is enough; all three are necessary today to produce the skill needed to meet the employment requirements of the coal industry. Routine manual labor will no longer do the job in coal mining. In the United States coal mining has been mechanized to a degree unsurpassed anywhere else in the world, and the coal worker must be thoroughly

Jan 1, 1973

By A. J. M. Ross

BACKFILLING of stopes and other underground openings in the Homestake mine with sand tailings was undertaken primarily to reduce surface subsidence, which was wrecking much of the surface plant and a substantial part of the town, and to make possible maintenance of open-ings into the top of pillar stopes as mining progressed. The progress of this type of filling has been rapid and the results very satisfactory. Sand is now used in filling timbered stopes, shrinkage stopes, and abandoned drifts and crosscuts. A permanent pipe-line system for transporting sand tailings into the mine by gravity has been installed. The close filling of all stopes has resulted in important benefits other than the two major ones. The ore bodies in the Homestake mine occur as replacements in one distinctive layer of the metamorphosed pre-Cambrian sediments along a group of minor folds that together form a complex anticline that pitches at 35° to 40° to the southeast. The ore bodies are extraordinary for their magnitude. Widths of over 300 ft. have been mined and continuous ore has been followed for over 1200 ft. along the strike on some of the upper levels. Mining has been by shrinkage stoping with square-set mining of crowns and pillars. In the standard plan, shrinkage stopes 60 ft. in the direction of strike and the full width of the ore have been taken up to within 25 ft. of the level above. After filling, the crown has been mined by square-set methods, as have also the 40-ft. pillars between shrinkage stopes. In former days, when the supply of ore seemed inexhaustible, the mining of pillars was allowed to lag far behind because of the high cost. Many pillars were rather badly fractured on the level and in some places between levels. Coarse waste was used for backfilling all stopes, but in pillars it became increasingly difficult to maintain haulageways for the transporta-tion of this fill over the top of the stopes as mining progressed. The cost of maintaining these crosscuts was excessive and in many pillars prac-tically impossible. Therefore it became the practice to fill pillar stopes by tapping the back and sides of the stopes.

Jan 1, 1939

By H. M. Chance

Probably no other prosperous mining district is so little known as the Black Hills. The name leads one to assume that the district is a rolling country, consisting of more or less insignificant hills. Nothing could be more untrue, as the Black Hills is an elevated region rising to heights of 7000 and 8000 feet above sea-level, and broken by deep gorges and ravines. It is, in fact, a mountainous country in all that that term implies ; but the cañons and gorges are of such a character that the country is easily penetrated by railroad-lines, and good wdgon-roads exist to-day, affording access to every portion of it. Aside from the stampede to the Black Hills country in 1876, when gold-placers were discovered, and 15,000 or 20,000 people became infected with the mining fever, the region has never had a mining " boom." Quartz-mining commenced in 1877 or 1878, and has been successfully prosecuted ever since; the output of gold steadily growing, until at present the yield is about $7,000,000 annually. This great production is about equally divided between two entirely different classes of deposits. The first class, comprising the so-called quartz or free-milling ores (of which the Homestake mine is the principal producer), are veins of generally steep dip, in the older metamorphosed rocks, and similar to the veins worked in other mining districts. The second class of mines is located in what is known as the " siliceous gold-belt." The Siliceous Gold-Belt. This belt, more particularly to be described in this paper, lies in the immediate vicinity of Deadwood and Lead City, and covers an area 6 or 7 miles in length by 3 or 4 miles in width. The formation in which the siliceous ores are found is a nearly horizontal series of sandstones and shales, generally

Jan 1, 1901

By J. S. Aronofsky, J. P. Heller

This paper presents a mathematical analysis of the fluid mixing which occurs during flow through porous media. The analysis is based on the well-known diffusion equation with mass transfer term. It is pointed out that the use of this equation is justified by a single general assumption which does not specify any particular mechanism for the mixing process. Formulas are given for tracer fluid concentrations for two different boundary and initial conditions. Calculated numerical values compare closely with some published results of experiments in which no viscosity or density difference existed between displaced and displacing fluids. INTRODUCTION A phenomenological theory of the mixing and diffusion process is presented for the flow in a porous medium of two miscible liquids of equal viscosity and density. This is based on the classical diffusion equation which has been used to explain such processes as Brownian motion and heat conduction. More recently, this same differential equation has been derived (in a manner analogous to the work in this paper) for the problem of heat transfer during fluid flow in porous media.' The purpose of this paper is to show that the available experimental evidence justifies the use of this simple diffusion equation. The recent data of Koch and Slobod' and of von Rosenberg' are used to test this diffusion model by comparing concentration curve shapes and calculating "effective diffusion coefficients". The same differential equation for miscible displacement, as well as some experimental confirmation, has been published in Japan by K. Yuhara,4 and recently in this country by Day.5 Yuhara arrives at his con-clusions through analogy with turbulent diffusion, de- scribing the microscopic velocities in terms of "eddy motion" — although unlike the ordinary eddies of hydraulics. Day proceeds directly from Scheidegger's6 concept of the dispersion of velocities. After a very clear description of Scheidegger's theory, he applies it to calculate the motion first of a drop and then of a thin layer of brine into a bed of sand. For the latter case, Day derives and solves the same differential equation as Eq. 1. In our presentation no detailed assumptions are made as to the mechanism of displacement. In fact, several microscopic processes may well be involved—the only general assumption made is that their cumulative microscopic effect may be described by the addition of a second-order term to the differentirl equation of continuity. DERIVATION OF EQUATIONS Let us consider the displacement in one spatial dimension of Liquid A from a sample of porous media by Liquid B. A and B are assumed completely miscible and of the same viscosity and density. Both fluids are also assumed to be incompressible and it is specified that the total flow is the same for any cross section. It is evident that the rate at which Liquid B would be carried across such a section would be given by the product of the fluid velocity, u, and the concentration, C, of B in the fluid mixture assuming the constancy of u and C over the cross section. Since the volume of Fluid B is conserved, this leads to the one-way wave equation - u ?C/?x = ? ?C/?l where x increases in the direction of u—the "Darcy velocity", Q/A (or velocity flux per unit area), and 9 is the fractional porosity. Consider the case where fluid B is displacing Fluid A and the concentration distribution is represented by a step funcLon (with a vertical discontinuity). The solution of Eq. 1 predicts no mixing

Jan 1, 1958

By H. I. Smith

WHEN Spain established colonies on the North American continent, some of her land grants, in what is now the United States, reserved to the Crown deposits of gold, silver, and mercury. Later mineral rights were reserved under some of the English Crown charters involving land in the eastern part of North America, with provisions for the payment of royalties thereon to the Crown. Reservation of mineral deposits to the United States Government was instituted by an ordinance of Congress on May 20, 1785, which applied to such deposits in the Northwest Territory, then north of the Ohio River and east of the Mississippi River, and provided that there should be reserved "one third of all gold, silver, lead, and copper mines, to be sold or otherwise disposed of as Congress shall hereafter direct." Little was known then of the mineral resources of the country; the Great Lakes copper region had just come into the possession of the United States by treaty and much of the western mineral land still belonged to France and Spain.

Jan 4, 1954

By Alfredo O. Del Campo

Copper concentrate smelting has been the subject of significant improvements during the last three decades. Several new processes are being successfully applied at industrial scale, and probably no new copper smelter will be built with the old-fashioned reverberatory furnace. However, because of financial reasons, the existing reverberatory furnaces, which smelt a significant part of the copper produced worldwide, are likely to last at least another decade. Some improvements have been introduced in reverberatory smelting practice, the best known being the use of oxyfuel burners. However, efforts to optimize the reverberatory furnace operation through automatic control have not been reported. The complex and semi-continuous nature of the process makes it dificult to find a suitable strategy for its automatic control. Many factors, like fuel economy, productivity, boiler capacity, refractory wear, converter slag return, copper losses in discard slag, etc. are interconnected. The chosen control strategy should take into account all these factors to achieve good economic results. A detailed discussion of these problems is presented.

Jan 1, 1984

By W. N. Logan

THE petroleum industry in Indiana made no extraordinary progress during the year 1924. The surplus stock of crude, brought about by the production of 732,407,000 bbl. in the United States in 1923, as against a production of 557,531,000 bbl. in 1922, had its effect on prices though there were slight advances. These advances were more than offset by the continued high prices of labor and materials. TRENTON OIL DEVELOPMENT The principal development of the year occurred in the Hotmire pool of the Ridgeville district in northeastern Indiana. This pool lies partly in the southern part of Jay County and partly in the northern part of Randolph County. The oil is obtained from the upper horizon of the Trenton. During the past year, about 30 wells were drilled, 21 of which were productive. An attempt to extend the Hotmire pool to the northeast resulted in the discovery of a gas-producing area near the center of Pike township in Jay County; 10 wells have been drilled 4 of which produced oil. The Harvey Orr No. 1 produced 900,000 cu. ft. of gas per day.

Jan 3, 1925

By E. R. Borcherdt

IT is notable that the first large-scale mine operation equipped entirely with detachable bits was the Badger State mine of the Anaconda Copper Mining Co. in Butte, Montana, just 30 years ago. This mine in 1922 was producing approximately 1200 tons of ore per day. Much of the data presented in C. L. Berrien's article1 describing the development and installation of the Hawkesworth detachable drill bit were obtained from these operations. As in any pioneering effort, no precedent existed and many difficult problems required solution, so that the changeover to detachable bits at all Butte hill mines was not completed for 6 years. There was widespread disbelief as to the probable efficiency of the new installation.

Jan 2, 1953

By T. F. Witherbee

The following calculation of heat requirement covers the working of the furnace from January 25th to February 14th, inclusive. A short time previous to the first date the furnace had been working rather badly, so that 300 lbs. of ore and 500 lbs. of limestone were taken off. Within the time mentioned the quality of the iron passed

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 Noel Jarrett, Allen S. Russell, Bernard M. Starner, Stanley C. Jacobs

A process has been developed to refine high-grade commercial aluminum to 99.99 pct purity. This enzploys precipitating titanium, vanadium, and zirconiu~ as borides. The upgraded liquid is partially crystallized under conditions that prevent massive freezing. The resulting crystals are compacted and washed free of adhernig impure liquid aluminum by melting the top crystals while draining off the liquid. Finally, the bottorn portion of the crystal bed is melted and removed as the pure product. Leakage from the crystallizat~on vessel is prevented by a layer of alumina granules that are not wetted by the aluminum. ALUMINUM of 99.99 pct purity is desired in a variety of applications. These include electrolytic capacitors, catalyst carriers, decorative trim for automobiles, aluminum powder, and high-strength alloys. This market was for many years filled by refining commercial-purity aluminum in a three-liquid-layer electrolytic cell of the type originally proposed by Hoopes. To meet the expanding need for lower-cost high-purity aluminum, a refining process based on chemical precipitation and fractional crystallization was developed. The availability of this super-purity metal was announced in November 1961, and three patents1-3 relating to the process were issued in 1965. An alternate process, comprising continuous fractional crystallization of a portion of a liquid-aluminum feed stream under conditions of strong agitation at the surface of the growing crystals, was published by Dewey4 in 1965. Earlier proposals by Larsen 5 and by Regner6 also emphasized the advantage of rapid agitation of the melt in the region of the crystallization zone. johnson 7 more recently proposed extracting purified aluminum from an alloy by mixing crystals of pure aluminum into the molten alloy, allowing them to grow, and removing the pure aluminum as by casting. Aamot's 8 recent patent describes a freeze refining process that involves controlled partial freezing against cooled moving retort walls. The Alcoa process to produce high-purity aluminum is schematically described in Fig. 1. Relatively pure aluminum is treated with boron in a furnace (a) to precipitate high melting point impurities, the supernatant liquid is transferred to a crystallization vessel (b), the liquid is partially crystallized with mechanized "tamping" (c), the remaining impure liquid is tapped out and the crystal bed partially and directionally re-melted to wash the underlying crystals (d), and the balance of the crystals is melted and withdrawn as high-purity aluminum (e). PRINCIPLES OF FRACTIONAL CRYSTALLIZATION Commonly encountered impurity elements form two types of systems with aluminum. Titanium, vanadium, and some other transition elements in periodic groups IV, V, and VI are the high melting point impurities. Initial solidification of aluminum containing small percentages of these elements will produce solid markedly enriched in them at temperatures slightly above the freezing point of high-purity aluminum. Most other elements, such as silicon and iron, produce eutectic type reactions with aluminum. In this case the first solid formed on initial solidification at temperatures slightly below the freezing point for high-purity aluminum is markedly depleted while the remaining liquid is enriched in these elements. The equilibrium change in composition during solidification of aluminum contaminated with silicon, iron, or titanium is illustrated in Fig. 2. Here, for clarity, the temperature scale is expanded beyond the

Jan 1, 1968

By A. L. Field, P. H. Royster

The first report on the slag viscosity work of the Bureau of Mines was made by one of the authors1 in 1916. It was concerned chiefly with the method of measurement. A paper2 on this phase of the work was presented to the Institute in February, 1917, and a similar paper3 was communicated to the Faraday Society. PuRpose and Scope of the Slag Viscosity Tables Following the development of a suitable method, the next step in the investigation was the measurement of slag viscosity over a wide range of temperature and composition, and the arrangement of this experimental data in the form of tables for the use of the blast-furnace operator. It is believed that this information, if used intelligently, should help the operator to reduce losses in production caused by off-grade pig iron, to improve fuel economy, to promote operating efficiency, and to extend present-day practice to meet the increasing need of smelting lean and complex ores. In the experimental part of the work it was carly, observed that the temperature-viscosity relations of a slag were determined, for practical purposes, by the relative proportions of the three major constituents, lime (CaO), alumina (Al2O3), and silica (SiO2), provided the percentage

Jan 1, 1918

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Jan 1, 1938