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By H. W. Rogers

The first steam shovels used in this country were built by the Otis Company, of Boston, about 50 years ago, but as they were of very crude construction and rather unsuccessful only a few were built. For possibly 10 years prior to 1884 successful steam shovels were made and used on certain classes of work, but it was not until that time that they were manufactured in quantities and began to play an important part in all classes of excavation. From that time up to the present day there have been gradual but continuous improvements on the original shovel, not only in the mechanical construction, but also in the design of boilers and engines which are best adapted to this class of service. In proposing a change from steam to electric operation we have to deal with a steam equipment which has not only proved its worth but has probably reached its highest stage of development and efficiency. That the electric shovel is a possibility cannot be denied, as at present from 12 to 18 shovels are in operation in this country. These shovels may be divided into three classes: the friction electric, which is operated by a single constant-speed motor with friction clutches; the three or four motor direct-current equipment, and the three or four motor alternating-current equipment. It is the second and third classes that I wish to deal with, as the first class does not compare favorably with the steam shovel so far as speed is concerned, although it may be operated as cheaply. There is probably no other class of machinery that presents a duty cycle as severe as that of the shovel, which is very short, varying from 7 to 12 sec. on the hoist, from 7 to 12 sec. on the thrust, and from 10 to 18 sec. on the swing, making a complete cycle in from 17 to 30 sec., and the motor to meet these requirements must have a sufficiently low armature inertia to permit of rapid acceleration and quick reversals under small power. It should also be a motor of rugged design, as it must be subjected to severe overloads and shocks and frequent reversals. This is especially true of the hoist motors and, to a lesser degree, of the swing motor; the thrust motor being practically stalled during the digging operation, although it may revolve or overhaul, according to conditions, and is operated at full speed only after the hoisting operation is completed.

Jan 1, 1915

By R. J. Arsenault

The yield drop that occurs in tantalum, Cu-AZ. and Ag-Al was investigated as a function of strain rate and at several temperatures. From the strain-rate dependence of the yield drop an activation volume was calculated and found to agree with that obtained from differential strain-mte tests. The temperature dependence of the yield drop can he explained in terms of the parameter temperature divided by activation volume. The yield drop is attributed to the rapid multiplication of dislocations during their motion. THE yield-point or yield-drop (upper yield stress minus the lower yield stress) phenomenon has been the subject of numerous investigations1-9 in both bcc and fcc metals. Also, there is no dearth of theoretical explanations10-20 to account for the yield-point phenomenon. The present explanations can be divided into two categories: the concept of dislocation locking and subsequent unlocking and the concept of multiplication of dislocations due to their motion. The Vottrell10,11mechanism of locking and unlocking, as applied to bcc metals, is the best known of the three proposed locking mechanisms. The two additional mechanisms, short-range order (SRO)7,16,17 and Suzuki-type locking,14,15 have been employed to account for the existence of yield points in fcc metals. The experimental evidence supporting the locking of dislocations, especially in bcc metals, is well-established. The dependence of the yield point on interstitial concentrations,' the kinetics of strain aging,2 and the observations by electron-transmission microscopy18 are some of these experimental results. The number of experimental observations supporting dislocation locking in fcc metals is not as extensive as in bcc metals. The unlocking of dislocations, which actually is supposed to account for the yield drop, is by no means well-established by experimental results for either bcc or fcc metals. The similar temperature dependence6 of the yielding and plastic flow is one of the experimental factors that cannot be explained in terms of the unlocking of dislocations. The mechanism based on the rapid multiplication of dislocations has been proposed by several investigators. 18-20 Johnston and Gilman19 have proposed a detailed model of yielding, which was later employed by Hahn21 to predict quantitatively the yield drop in iron. This mechanism was discussed in a previous paper22 and several limitations were pointed out. In the same paper it was shown that an activation volume could be obtained from the strain-rate dependence of the yield drop and that a correlation existed between the activation volumes obtained from the yield-drop tests and from differen-

Jan 1, 1964

By W. A. Alexander, L. M. Pidgeon

Metallic magnesium and similar meta!s near the top of the electromotive series have been commercially produced by the electrolysis of a suitable molten salt. Despite the success of electrolysis, sufficient inherent objections exist to justify the search for a direct reduction method. Such a method would widen the choice of raw materials, relax the rigid electrolytic requirements of raw-material purity, and, by obviating direct current, over simplicity of plant equipment. This paper describes the pilot-plant development of one such method, which has subsequently formed the basis of operation of six commercial plants-—one in Canada and five in the United States. Unlike aluminum, magnesium is volatile at relatively low temperatures. This physical property, which produces its greater inflammability, also offers possibilities of direct reduction that are not available with aluminum. In the basic reaction MgO + X = XO + Mg if X is nonvolatile, the reaction may be forced to the right at high temperatures by the evolution of magnesium vapor, despite the large negative value of the heat of reaction which is almost bound to follow the use of any available commercial reducing agent. The cheapest X, of course, is carbon, but this notable advantage is modified by the production of a volatile oxide CO, so that both Mg and CO are evolved simultaneously. Elaborate devices are required to shock-cool the equilibrium mixture in order to prevent back reaction. At best, a pyrophoric powder is produced, requiring redistillation to provide marketable metal. The practical development of this process, is due to F. Hansgirg, and his methods have recently been given a trial on a grand scale in the plant at Permanente, California, and smaller plants were previously built in Austria, England and Japan. A second class of reducing agents is available, those which produce nonvolatile oxides. When XO and X are relatively nonvolatile, magnesium is the only volatile member of the system. It may be evolved in a simple distillation step and condensed without the formation of powder. Two basic requirements must bc fulfilled in order that such a reaction may proceed: I. The reducing agent when heated to a reasonable temperature in presence of MgO must cause the production of an appreciable equilibrium vapor pressure of magnesium. This magnesium must be removed continuously. z. A supply of heat must be maintained. In order to fulfill the first requirement, operations must be conducted in vacuo or in a stream of Hz. Owing to its greater efficacy and relative safety, vacuum has been employed in the work to be described. The choice of a suitable reducing agent formed the subject of an extensive series

Jan 1, 1944

By W. W. Anderson

WEBSTER'S dictionary gives the following definition for "efficiency": "Effective operation as measured by a comparison of actual and possible results." Engineers think of this definition in terms of the ratio of output to input, and since they learn early in their careers that output cannot be greater than input, they usually consider a machine to be 100 pct efficient if it operates perfectly, and to be 0 pct efficient if it performs no function at all. Hence, values of efficiency between 0 and 100 pct reflect how well or how poorly a machine has performed; and any method of calculation for determining efficiency that does not show 0 pct when no function has been performed, and 100 pct when a perfect job has been done, is looked upon with skepticism. The literature on performance of coal cleaning equipment contains many references to efficiency of performance, but there is no unanimous agreement among the authors as to how the efficiency should be determined. Numerous methods of calculation have been proposed from time to time, most of which attempt to fulfill the definition of efficiency by comparing output to input, using float and sink data or some other characteristic, such as dry ash content, as a measure. However, to the best of our knowledge, none of the methods advocated in the past meets the requirements of a practical engineer in judging cleaning plant performance, because the formulas previously advocated do not confine numerical values of efficiency between 0 and 100 pct, and the literature does not contain an explanation of the meaning of the formulas under various conditions of operation. Therefore, the following formula, based on Yancey's concept of misplaced material,' is proposed as a measure of efficiency which will conform with the requirement that the values of efficiency must be confined between 0 and 100 pct: Quantitative Efficiency = 100.0 — (pct wt sink* in cleaned coal + pct wt float* in refuse)t • At gravity of separation effected by cleaning equipment. t Total product basis of the size of material under study. It will be noted from this formula that an actual separation between cleaned coal and refuse must be effected by the cleaning equipment before the formula can be used; also, that improperly distributed material in the products must be determined at the gravity of separation, rather than at some arbitrary test gravity. It will further be noted from this formula that the weight of cleaned coal recovered and the weight of refuse rejected by the equipment have their proper influence, because the sink in the cleaned coal and the float in the refuse are expressed in the equation as a per cent weight of the total products of a definite size discharged from the plant. This formula has one other important characteristic, in that it requires no knowledge of the feed to the cleaning unit. Comparison of data on the products from a cleaning unit with the feed to the unit can lead to many worthwhile conclusions. but not in regard to the efficiency of separation effected by the equipment. The reason for this is that mechanical fracture of many pieces of the feed occurs during the cleaning process; and in some processes crushers are deliberately incorporated in the flow to produce size degradation to free impurities from cleaned coal, thereby altering the washability characteristics of the material while in process. When such size degradation occurs in a cleaning process, it is quite possible that this crushing will result in a reduction in ash content in the finer sizes of cleaned coal when compared with the raw coal feed, even though no true separation has been effected. This is due to the fact that the finer sizes have been "sweetened" by the crushing of cleaned coal from some coarser size. Many formulas advocated in the past show high cleaning efficiencies

Jan 1, 1951

By P. de Vitry, Oliver Bowles

The conventional wire saw, introduced in the slate district of Pennsylvania by the Bureau of Mines in 1927, and used thereafter with remarkable success, consists of a three-strand steel cable having a diameter of 93/16 or 54 in. The equipment, methods used, and accomplishments have been described in previous reports.'-I Similar wire saws have been used extensively in Europe, in quarries and in shaping blocks and slabs. During the period 1932-1937, competition in European marble quarries was keen, prices were declining, and orders were not abundant. When orders were received, there was generally an urgent call for speedy delivery. Because of the diversification of demands, it was impossible to maintain in stock marbles of the color, size, and thickness called for, therefore the material generally had to be fabricated after the order was received. The wire saw enabled stone mills to fill such orders promptly. With skillful handling, 7/8-in. slabs could be cut accurately and rapidly with this equipment. It was found, however, that, as the three-strand wire became worn, the cutting speed dropped rapidly and the cut became less accurate. To overcome these difficulties the junior author of this paper, after much experimentation, developed a new type of wire consisting of a single ribbon-shaped strand, twisted as shown in Fig. I. To ensure straight sawing, the twist of the ribbon is reversed from right to left and vice versa every 25 ft. The reversal of the twist is shown in Fig. 2. This is not a new feature; it has been used to advantage with the three-strand wire both in Europe and America for several years. However, some difficulties arise in splicing the three-strand reverse-twist wire, and the claim is made that cutting is less accurate as the wire becomes worn. To make an endless belt, the ends of the single-strand wire are brazed together with Pyrophyllite is a hydrous aluminum silicate (A1²Si4O10(OH)²)1 that, occurs in both the foliated and the massive forms. The foliated variety resembles talc in that it has a greasy feel, a pearly luster, perfect basal cleavage, and usually is white, although in some deposits apple green, gray, brown, russet, and nearly black specimens may be found. The massive deposits of pyrophyllite yield pearl gray to light tan aggregates, some of which are so friable that they may be crushed with the hands into a fine gritty powder. Frequently specimens of the massive variety are studded with crystals of radial structure. The mineral has a specific gravity of about 2.7 and a hardness of 1 on Mohs' scale.' The chemical composition of commercial pyrophyllite reflects the mineralogy of the deposits, which, according to Stuckey12 contain pyrophyllite and quartz, together with chloritoid, sericite, and a few other minerals of negligible importance. The latter, he says, ('are noticed in but small quantities to the extent they might occur as accessory constituents of an igneous rock or as products of regional metamorphism or weathering. . . . Quartz is abundant everywhere except in the very best grades of pyrophyllite." Analyses of specimens from several deposits are listed in Table 1, showing that the silica content varies from 57 to 73 per cent and the alumina from 22 to 33 per cent. Oxides of iron and the alkaline metals are collectively present, with one exception, to an extent of about 1 to 1.5 per cent. Sources At present, North Carolina is the only commercial source of pyrophyllite, although the mineral has been found in small quantities in many parts of the world. The North Carolina deposits are in the "Carolina Slate Belt," which traverses the center of the state in a southwesterly direction. This belt varies from 8 to 50 miles in width. Two major deposits of pyrophyllite lie along its eastern boundary, in what is known as the Deep

Jan 1, 1942

By G. W. Thomas

A simplified mathematical model of underground conduction heating in a system of limited permeability is presented. The model applies to underground retorting of oil shale, or to reservoirs containing extremely heavy oils. We assume that heat is introduced at a constant rate into a horizontal fracture which communicates between wells. The radial temperature distribution along the fractured surface is approximated by a step-function. Heat transfer away from the fracture is assumed to be by vertical conduction, and all convection effects are neglected. The model also takes into account the possible temperature dependence of thermal conductivity. A general expression for calculating the growth of the step-function temperature distribution with time is derived. The use of this expression and solutions to the one-dimensional heat equation make it possible to construct isotherms. Expressions for calculating .oil recovery, well spacing and heat efficiency are also given. An example calculation is presented for the conduction heating process in oil shale. Finally, the effect of the heat transfer coefficient between the gas and the fracture boundaries is investigated. INTRODUCTION Thermal recovery processes of oil recovery fall into the four general areas of hot fluid injection, forward combustion, reverse combustion and conduction heating. The first three of* these processes have been rather extensively studied in the past decade from both the experimental and theoretical points of view. As a result, it is possible to make reasonable engineering predictions and analyses of these processes. Little attention has been devoted, however, to the conduction heating process other than to note its possible utility.I To a certain extent, conduction heating cannot be divorced from the other regimes cited above, insofar as these provide the source of heat energy. In the conduction heating process, heat is introduced (either by combustion in the forward or reverse mode or by hot fluid injection) into a small fraction of the total reservoir thickness. This fraction may be either a streak of high permeability or an interwell fracture. Heat penetrates by conduction into the adjacent, less permeable regions of the oil-bearing rock, where the direction of conduction is essentially perpendicular to the streak or fracture. The heated product then drains by gravity or is gas driven to production wells. Conduction heating is probably most applicable to systems containing immobile bitumens such as tar sands and oil shale deposits and perhaps to low-permeability reservoirs containing highly viscous crudes. The mechanism also acts in combination with other thermal processes where fingering or overriding of a bed occurs. It seems probable that in at least one field test, conduction heating of this type was very influential.2 In this presentation we give a first approximation to some of the quantitative aspects of conduction heating. Marx and Langenheim3 treated a similar problem where they focused attention upon the injection interval, which spanned the total reservoir thickness. In their model, conduction heat losses to the bounding media imposed a practical limit on the calculated heated area. In the present study, however, we shall confine the injection interval to a small fraction of the reservoir thickness and assume it has no heat capacity. We therefore direct our attention to regions outside the injection interval into which the conduction of heat is beneficial. In particular, we will endeavor to locate specific isotherms in the media bounding the injection interval. Furthermore, we will construct our model to allow the thermal conductivity to vary arbitrarily with temperature. Thus the model will be applicable to underground retorting of shale where variations in thermal conductivity may be important. THEORY We consider the rock matrix to have no or, at best, a very limited permeability; consequently a

Jan 1, 1965

By P. de Vitry, Oliver Bowles

The conventional wire saw, introduced in the slate district of Pennsylvania by the Bureau of Mines in 1927, and used thereafter with remarkable success, consists of a three-strand steel cable having a diameter of 93/16 or 54 in. The equipment, methods used, and accomplishments have been described in previous reports.'-I Similar wire saws have been used extensively in Europe, in quarries and in shaping blocks and slabs. During the period 1932-1937, competition in European marble quarries was keen, prices were declining, and orders were not abundant. When orders were received, there was generally an urgent call for speedy delivery. Because of the diversification of demands, it was impossible to maintain in stock marbles of the color, size, and thickness called for, therefore the material generally had to be fabricated after the order was received. The wire saw enabled stone mills to fill such orders promptly. With skillful handling, 7/8-in. slabs could be cut accurately and rapidly with this equipment. It was found, however, that, as the three-strand wire became worn, the cutting speed dropped rapidly and the cut became less accurate. To overcome these difficulties the junior author of this paper, after much experimentation, developed a new type of wire consisting of a single ribbon-shaped strand, twisted as shown in Fig. I. To ensure straight sawing, the twist of the ribbon is reversed from right to left and vice versa every 25 ft. The reversal of the twist is shown in Fig. 2. This is not a new feature; it has been used to advantage with the three-strand wire both in Europe and America for several years. However, some difficulties arise in splicing the three-strand reverse-twist wire, and the claim is made that cutting is less accurate as the wire becomes worn. To make an endless belt, the ends of the single-strand wire are brazed together with Pyrophyllite is a hydrous aluminum silicate (A1²Si4O10(OH)²)1 that, occurs in both the foliated and the massive forms. The foliated variety resembles talc in that it has a greasy feel, a pearly luster, perfect basal cleavage, and usually is white, although in some deposits apple green, gray, brown, russet, and nearly black specimens may be found. The massive deposits of pyrophyllite yield pearl gray to light tan aggregates, some of which are so friable that they may be crushed with the hands into a fine gritty powder. Frequently specimens of the massive variety are studded with crystals of radial structure. The mineral has a specific gravity of about 2.7 and a hardness of 1 on Mohs' scale.' The chemical composition of commercial pyrophyllite reflects the mineralogy of the deposits, which, according to Stuckey12 contain pyrophyllite and quartz, together with chloritoid, sericite, and a few other minerals of negligible importance. The latter, he says, ('are noticed in but small quantities to the extent they might occur as accessory constituents of an igneous rock or as products of regional metamorphism or weathering. . . . Quartz is abundant everywhere except in the very best grades of pyrophyllite." Analyses of specimens from several deposits are listed in Table 1, showing that the silica content varies from 57 to 73 per cent and the alumina from 22 to 33 per cent. Oxides of iron and the alkaline metals are collectively present, with one exception, to an extent of about 1 to 1.5 per cent. Sources At present, North Carolina is the only commercial source of pyrophyllite, although the mineral has been found in small quantities in many parts of the world. The North Carolina deposits are in the "Carolina Slate Belt," which traverses the center of the state in a southwesterly direction. This belt varies from 8 to 50 miles in width. Two major deposits of pyrophyllite lie along its eastern boundary, in what is known as the Deep

Jan 1, 1942

By Pol Duwez

The effect of the rate of cooling on titanium, zirconium, and thallium has been measured. For titanium-molybdenum alloys, it has been shown that for molybdenum concentration up to 8 pct the rate of cooling has no effect on the transformation temperature, but this temperature decreases progressively with increasing molybdenum content. HIGH strength to weight ratio and remarkable corrosion resistance have often been advanced as the main reasons for the present widespread interest in titanium. For the physical metallurgist, the existence of an allotropic transformation in titanium is another important factor. As in the case of iron, an allotropic transformation may be the fundamental property upon which the response of Ti-base alloys to heat treatment will depend. It is therefore important to study the mechanism of the allotropic transformation in pure titanium and the effect of alloying elements on this transformation. The existence of an allotropic transformation in titanium was reported in 1931 by Schulze.' Later, De Boer et al.' determined the transformation temperature from a discontinuity in the electrical resistivity at 882"42O°C. According to recent work," the transformation temperature is 882.5"&1°C. The crystal structure of the room temperature a phase is hexagonal close-packed and the high temperature fi phase is body-centered cubic.' It is well known that the high temperature /3 phase of pure titanium cannot be retained by quenching to room temperature. Several investigators'-' have shown that in alloys of titanium with metals which are soluble in the @ phase (for example, nickel, molybdenum, chromium, or manganese), the /3 structure can be retained, providing the concentration of the alloying element is greater than a certain critical value. Below this critical value, the alloys transform during quenching, at least partially, into a supersaturated solid solution having the a. structure. This behavior is very similar to that found in some iron binary alloys (Fe-Ni and Fe-Mh, for example) and is an indication that the transformation might occur by means of a shear mechanism analogous to that observed in the austenite-martensite reaction. The present investigation is mostly concerned with the effect of the rate of cooling on the /3 to a transformation. In the first portion of this paper, the results obtained on pure titanium are described. Two other pure metals, zirconium and thallium, were also investigated, because they both have an allotropic transformation like titanium, from an hexagonal close-packed to a body-centered cubic structure. For purposes of comparison, the effect of rate of cooling of the y to a transformation of iron was included in the experimental program. The second part of this paper deals with Ti-Mo alloys. The equilibrium phase boundaries involving the solid state reactions on the Ti-rich side of the system were determined and a study was made of the effect of the rate of cooling on the transformation of the high temperature /3 solid solution. Experimental Procedure Since the main object of the present investigation was the determination of the effect of the rate of cooling on a transformation temperature, the best technique available was that used by Greninger" in his study of the martensite arrest in carbon steels. This method consists of heating a very small sample of metal to which fine thermocouple wires are .attached, cooling the sample rapidly by means of a blast of helium, and recording the temperature vs. time curve on an oscillograph. The apparatus used in the present experiments was very similar to that described by Greninger, except that the heating coil was made of molybdenum wire (0.040 in. in diam) and the assembly was placed in a bell jar under vacuum of 10" mm Hg or less. The specimen con-

Jan 1, 1952

By Fernando A. Buttazzoni, Jorge C. Munita

INTRODUCTION To provide a background for discussing long-range financial and operational planning in the mining industry, we would like to begin by describing the role of planning in Codelco. Having taken part in Codelco's planning process since its inception in 1976, we believe that useful conclusions may be drawn for the experience we shall relate. Although Codelco is a state-owned copper producer with an important role in the Chilean economy, and is consequently very much at the core of national economic planning, the following discussion is not related to government planning but to planning in an operationally and commercially autonomous business concern with all the characteristics of a private enterprise. The only way in which Codelco differs from private business is that its financial successes or failures have a more direct impact on the countryls ability to participate in international trade, to meet its international financial commitments and to achieve economic and social growth. Thus, what follows is a presentation of the conceptual features of the planning process at Codelco and a description of the current state of planning in the company. It is understood that certain conceptual components are basic to any formal strategic planning process, and that the components of Codelco's planning process are therefore generally applicable in similar companies. First, the mission and scope of operation of the company must be defined, and goals must be set to achieve leadership in the business activities in which it engages. Having set this framework, the company must constantly scrutinize those internal elements and conditions that are essential to fulfill its mission and achieve leadership. In Codelco this means considering the actions which can best serve to define future capacities, taking into account declining ore grades and determining optimal production levels; watching over the productivity and skills of its labor force, and; keeping abreast of technological developments in non-ferrous mining and processing. If the company is proficient in these three areas, it will be cost efficient and there- fore competitive. Codelco must also watch over the business environment in which it operates. The company sells concentrate, blister and fire-refined copper, wirebar, cathodes and wire rod and must study the behavior of the market to attune its product mix to the demand existing for these various commodities in the international market and the copper consuming industries in general. By examining the internal and external environment in which the company conducts its production and commercial operations, it is able to diagnose its strengths and weaknesses, to identify the hazards it must overcome to stay ahead in the copper trade, and to take full advantage of business opportunities as they arise. CHARACTERISTICS OF CODELCO-CHILE Although our subject is Codelco, readers should keep in mind that there are many other copper producing operations in Chile, all in the private sector, and many valuable deposits that may be exploited in the future. Codelco is the only public sector copper producer, although there is one other state enterprise that operates smelting and refining facilities principally to serve a large number of very small producers of copper concentrate. In the mid-19801s, Codelco is the largest copper mining company of the world and the largest producer and exporter of copper with yearly sales totalling close to US$ 1.8 billion in 1983. Moreover, it is a highly profitable copper company at a time when the world copper industry as a whole has reported losses for several years. Codelco-Chile is a fairly new organization. The history of the four mining operations it manages is much longer: El Teniente was opened in 1912, Chuquicamata in 1915, Salavador in 1959 -- with installations of the old Potrerillos mine dating from 1927 -- and Andina in 1970. Having had separate owners, the management systems of the

Jan 1, 1985

By T. J. Neil, C. B. Manula, O&apos

This paper describes a simulation model developed to analyze truck haulage problems as they are found in the open pit mining and construction industries. The model, using a computer, cycles trucks between their assigned loading and discharge points over a measured haulage route. Required parameters for model operation include the number and type of operating trucks and loading units, maximum truck acceleration and velocity, a realistic vehicular deceleration rate, and the ratio of ore to total material loaded at each origin. Equipment performance characteristics, system profile characteristics, and service time distributions are also included. Records are kept of all waiting times at the loading and discharge points and of any truck interference on the haulroads (moving queues), and a current journal of ore and waste production is maintained. In the pursuit of reduced truck haulage costs, management is confronted with three particular problems in truck and shovel mines. 1) What is the best assignment of trucks to shovels in the system? 2) If new equipment is needed, which new shovels and trucks will operate most effectively in the system? 3) Are any physical changes in the haulage layout warranted to increase productivity? For small open-pit mines the answers to these questions may be obvious. However, for larger operations complex problems are encountered with production and system configuration relationships usually being unknown. The mining simulator discussed here was developed to aid management in evaluating alternative pit-haulage schemes to reduce the risk involved in the selection and assignment of equipment in open-pit truck and shovel mining systems. There are two fundamental differences between this model and other truck haulage simulators: one is a difference of degree, the other a difference of kind. This model can simulate more complex systems than previous models by allowing transportation of material from multiple mine faces to multiple mine destinations and by being readily adaptable to a wide variety of mine layouts. The second difference that characterizes the model is the standard simulation of truck movements. This technique enables each vehicle to perform according to its mechanical capabilities and the physical profile of the haul road. Among the similarities between other truck haulage simulators and this model are the recognition of the probabilistic nature of certain service times in the system and the use of a digital computer to operate the model. The computer inherits the staggering computational and bookkeeping work loads and provides a means for compressing the time scale by allowing several days' operations to be simulated in a matter of minutes. DISTRIBUTION SYSTEM Basic aspects of the system analyzed here may be brought into focus by referring to Fig. 1, a schematic of the typical open pit truck-shovel operation programmed into the simulation model. A single stage materials handling scheme is employed to transport both ore and waste. Loaded trucks are directed to a crusher or stockpile if ore is being mined, while overburden is conveyed to a waste dump. Each shovel has a given number of assigned trucks and has two associated haulage routes, one for waste and one for ore. Up to 5 shovels and as many as 10 trucks per shovel (or a total of 50 trucks) can be operated in the model. Essentially the model, using a computer, cycles trucks between their assigned shovels and the proper discharge points over measured haulage routes. Required input parameters for model operation include the number of shifts to be simulated, operating time per shift, number and type of operating trucks and shovels, maximum truck acceleration and velocity, a realistic vehicular deceleration rate, and the ratio of ore to total material loaded at each shovel. Equipment performance characteristics, system profile characteristics, and service time distributions are

Jan 1, 1968

By Donald L. Johnson

REVIEWS on the applications of optical ellipsometry to the study of polycrystalline metal surfaces and oxidation-corrosion processes have been presented by Kruger and Hayfield, and white.' In view of the ability of this method to measure differences in film thicknesses to below 10A, ellipsometry was used to study nitrogen chromium surface reactions during the early stages of growth. Chromium specimens in the form of a vapor-deposited film on glass substrate were exposed to purified nitrogen for selected time increments at 700°F. This temperature was selected because it gave the best combination of exposure time and thickness change. Ellipsometer readings were taken immediately after exposure to determine polarizer and analyzer angles using monochromatic light at 5000A.2 After each set of readings was obtained, the specimens were exposed for an additional increment of time in nitrogen and a new set of readings taken. Film thickness was computed from LUCy's3 equation for nonabsorbing thin films. In Lucy's derivation, two equations for computing thickness were obtained. Peterson and Bashara4 have discussed the validity for using one of these equations when an independent value of the refractive index is used. The film was assumed to be Cr2N (Ref. 5) and was determined to be optically isotropic. Optical isotropy was determined by rotating the films about the surface normal and noting that the azimuth and phase difference did not change. Since Cr2N is reported to be hexagonal, it is assumed that optical isotropy results from randomly oriented grains of very small size. The thickness was computed using refractive indexes varying from 1.6 to 2.0. The maximum variation in computed thickness over this range for a given exposure was 20 pct. The data reported are based on a refractive index of 1.8. Reference readings on the metal surface were taken just prior to nitrogen exposure. No attempt was made to deoxidize or denitride the metal surface after vapor deposition, although specimens were observed to gain approximately 2A per day in air at room temperature. Twelve experimental exposures were made on a single sample over a total accumulative time of 132 min. Additional experiments over longer time intervals were made to determine the effect of heat up and cool down during and after exposure. As a result of these experiments, the thickness computed for each exposure was increased by 4A. Time profiles based on the linear law are similar to the oxidation profiles for many metals.' The rate of accumulation of the nitride film is rapid at first and slows down as the film thickness increases. The parabolic law x2 =k1t +k2 [11 gives a straight line in Fig. 1; therefore, it is concluded that the kinetic behavior is parabolic. Sey-bolt and Haman5 reported similar results for CrlN on chromium at 1000°C although the order of the thickness was much greater (10-3 cm). Analysis of thickness data based on other values of refractive index shows that the curves are shifted vertically but retain the same shape. According to Mott,7 it is expected that for thin films k2 is zero in Eq. [I]. Since the plot in Fig. 1 is linear through the origin, k2 is zero and the parabolic nature of the film growth is further indicated. The parabolic rate constant, calculated from the slope in Fig. 1, is 4.4 x 10-16 sq cm per sec. The corresponding work (+) required to bring an electron from the metal into the oxide is estimated to be 1.5 ev.7 The author thanks professor N. M. Bashara and D. W. Peterson, Electrical Materials Laboratory, University of Nebraska, for making the ellipsometer available and providing technical assistance. The Electrical Materials Laboratory is supported by the office of Naval Research, NONR 3718 (01). This work was partially supported by the Engineering Experiment Station and the Mechanical Engineering Department, University of Nebraska.

Jan 1, 1967

By K. D. Jacob, W. L. Hill

Fluorine recovery in the United States has been restricted chiefly to manufacture of ordinary superphosphate and wet-process phosphoric acid. However, there is an expanding use of fluorine by industry in the form of hydrofluoric acid. Improved methods for recovering fluorine from phosphate rock would yield an adequate supply of fluosilicates and a substantial contribution to the hydrofluoric acid requirement. THE bulk of natural phosphates is comprised of calcium phosphates, which are usually apatites;' calcium aluminum phosphates such as pseudowavel-1ite; and aluminum phosphates, which occur in extensive deposits, notably on Grand Connetable Island." Nearly all deposits of current commercial value are apatite, often with admixed material of other classes. In the trade coarsely crystalline material, mainly fluorapatite of igneous origin, moves as apatite, whereas the fine-grained material of sedimentary origin is marketed as phosphate rock. The latter classification includes the phosphate shales in western North America. Although several genetic varieties of phosphate rock are recognized, most known reserves are of marine origin. Calcium phosphate is concentrated from Florida land pebble and Florida hard rock deposits and from Tennessee brown rock deposits. Phosphate Rock as a Fluorine Carrier: More than two decades ago Reynolds and Jacob surveyed the fluorine content of commercial phosphate rock from various parts of the world. Later these samples were re-analyzed by a new and more reliable method of determining flourine. The revised results with interim additions of new samples, published a few years ago in a brief summary," are shown in Figs. 1-3. Through the points representing the respective varieties of rock medial straight lines are drawn. In some instances the lines are well defined by the points; in others, where scattering renders the position of the best line uncertain, the greater weight is given points near the end of the covered range. Wherever distinct varieties are not recognized among regional deposits the custom is to identify the sample with the geographical location of the deposit. Although all rock varieties cannot be classified with certainty, the plotted results exhibit three distinct types of rock in which the F to P2O5 ratio diminishes, remains sensibly constant, and increases, respectively, as the grade of rock increases. Existence of two types of Florida land pebble, Fig. 1, is supported by additional results recently supplied by a producer." Generally speaking, phosphate rock from continental deposits carries considerably more fluorine, and apatite and insular phosphates less, in proportion to the phosphorus content than is required by the formula for fluorapatite. Exceptions to this are apatite from Canada and certain samples of Florida waste-pond phosphate. The manner in which excess fluorine is held and the nature of the carbonate in carbonate apatite are discussed in recent articles. High-grade calcium phosphates with F to P2O5 ratios less than 0.05 have been found in Curacao, Christmas Island," Mexic,13 and elsewhere. The calcium aluminum phosphates and aluminum phosphates aften carry notable amounts of fluorine. The F to P2O, ratio for one sample of pseudowavel-lite concentrate was 0.05." Aluminum phosphate from Grand Connetable Island is very low in fluorine, whereas some aluminum phosphate minerals" carry as much fluorine as highly fluorinated phosphate rock. Fluorine in Domestic Phosphate: The curves shown in the figures permit interpolation of the F to P2O5 ratio corresponding to the midpoint of selected ranges in grade of any one of the several varieties of phosphate. Multiplication of the interpolated ratio by the mid-range grade (P2O5) yields a figure for the percentage of fluorine in this grade. Then, with the use of the mid-range grade, the amount of fluorine held in phosphate reserves can be estimated, see Table I. Accordingly, on Jan. 1, 1950, the fluorine complement of known domestic reserves exceeded 420,000,000 long tons of the element. A previous figure," deduced in 1940 by a simpler procedure

Jan 1, 1955

By Frank T. Weems

The basic operating properties of the DorrClone are discussed and certain metallurgical applications which exploit these properties are presented. An effective method of controlling the consistency of the DorrClone underflow is described. THIS paper presents in a general manner some of the applications of the DorrClone (Dutch States Mines cyclone) in the metallurgical and heavy chemical fields. Technical or mathematical analyses of the design factors and of the mechanism whereby the cyclone effects its separation have been capably treated in a series of previous papers.'-' Engineering and operational aspects of a few installations and laboratory tests have been chosen to emphasize the unique classification characteristics of the DorrClone and to illustrate its utility in certain difficult situations. Its use in coal washing and heavy-media operations are not considered. Some 25 or more industrial installations and applications of the DorrClone in the metallurgical fields are operating at present, ranging from single trial units to complete batteries composed of a multiplicity of units. The data were obtained in part from some of these, and in part from operations of units at the Westport Mill, the laboratories and testing plant of the Dorr Co. at Westport, Conn. The DorrClone was developed, in principle, by the research dept. of the State Mines, Limburg, Holland, during World War 11. Its first use was to dewater a —50 micron sand used as a dense medium in coal washing. The metallurgical wet cyclone was first publicized by Driessen and Yancey and Geer.6 ince then the cyclone has been studied as a means of de-watering, heavy media concentration, and fine size classification. The DorrClone is shown in Figs. 1 and 2. pulp to be treated is pumped through the feed nozzle, A, into the feed cylinder, B. The feed nozzle is of rectangular cross-section, the outermost edge being tangent to the feed cylinder walls. Feed shims, C, may be attached to the innermost walls to adjust the open area of the feed nozzle available to entering pulp. The tangential injection of feed under pressure supports the vortex action of the pulp that has been detained in the cyclone; the centrifugal forces of this vortex throw the coarse particles to the walls of the cone, D, where they collect and pass downward and out through the adjustable apex opening, E. The fine particles contained in most of the water that entered with the feed pulp move to the inner spiral of the vortex and flow out through the vortex finder, F, contained in the orifice plate, G, which forms the top of cyclone chamber. The overflow pulp may then spray freely into a spray chamber, H, surmounting the cyclone or pass directly into a 90' elbow, I, of somewhat larger diameter than the vortex finder. Installation Where more than one cyclone is to be installed, best overall efficiency and ease of control are obtained with equal distribution of both pulp volume and weight of solids among the cyclones, and with the same pressure being maintained at each cyclone feed nozzle. Fig. 3 is a suggested arrangement for feeding two or more cyclones in parallel. Feed pressure control: Under normal operating conditions it is not unusual for some surge to exist in the feed that is to be sent to the cyclone. As will be discussed later, surges in the solids content of the pulp are usually not of great operational consequence, but surges in feed volume are a problem. Several systems have been devised for absorbing surges in feed volume, each involving a surge tank. If the surges are small and fairly rhythmical, a tank of sufficient size might be installed to dampen out

Jan 1, 1952

By R. L. Pober, L. Kaufman, E. V. Clougherty

Composite structure of hafnium, zirconium, and titanium diboride with additions of metals and/or compound phases were prepared by reactive high-pressure hot pvessing and evaluated in air and in mixtures of oxygen and helium at temperatures up to 2400°K and at linear flow rates of 0.1 to 0.2 ft per sec stp. Oxidation characteristics, principally depth of conversion of diboride to dioxide, were determined by a gas analysis method and by postoxidation metallo-graphic analyses. The best oxidation resistance was observed for hafnium diboride-silicon carbide composite which exhibited a 6-mil conversion to dioxide for a 1-hv oxidation in air at 2200°K compared to a 20-mil recession for hafnium diboride with no additive. Analogous improvements in oxidation resistance were observed for the addition of silicon carbide to zirconium. EARLIER investigationof the oxidation behavior of the diborides of the transition elements of Groups IVA and VA in flowing gas mixtures have produced a ranking of these compounds according to which HfB2 is the most oxidation-resistant followed in turn by ZrB2 > TiB2 > TaBz > NbB2. Other ~tudies~-~ have examined the physical, thermal, and thermodynamic properties of these diborides to provide a sound basis for the more detailed and specialized investigations which are required to generate the information needed to assess the usefulness of such materials in high-temperature oxidizing environments. Recent oxidation studies7 have examined the effect of variations of metal and boron content on the oxidation characteristics of zirconium and hafnium diboride in flowing gas mixtures including air and helium plus oxygen. Results obtained in the latter investigation confirmed the superior oxidation resistance of HfB2 over ZrB2. Moreover, metal-rich compositions of hafnium diboride showed improved oxidation resistance over boron-rich compositions at temperatures up to 2000°K; at higher temperatures the differences are not distinguishable. In particular HfBI.7 exhibited a diboride to dioxide conversion depth of 4 mils in 1 hr at 2000°K; HfB2. 2 exhibited a conversion depth of 16 mils for the same exposure.~ Specimens for the study of metal and boron stoichiometry on the oxidation characteristics of HfB2 and ZrB2 were prepared by high-pressure hot pressing, a specialized type of fabrication which produced dense crack-free bodies of the desired compositions with no significant impurity introductions. The purpose of the present investigation was to explore the possibility of further improving the oxidation resistance of the metal-rich ZrB, and HfB2. The plan to accomplish this objective was based in the formulation of several diboride compositions containing other compounds and phases which are generally known to have good oxidation resistance in one or more types of test environment. Ideally, such additives would also enhance or at least not diminish other boride property values principally mechanical strength and those parameters related to thermal stress resistance. The additions selected for HfBz included zirconium, aluminum, and (Ta + TaB2) designed to produce ternary metal diborides, hafnium to form a discrete HfB phase, chromium to provide a metal skeletal phase, (Hf + Si) to form a discrete hafnium silicide phase, and Sic to provide a discrete second phase. Additions to ZrB2 included A1 + B designed to produce a ternary metal diboride, Zr + Si to form a discrete zirconium silicide phase, and Sic to provide a discrete second phase. All the specimens were prepared by high-pressure hot pressing which was used to advantage to fabricate billets of the variety of compositions desired and suitable for oxidation testing. The present investigation also included the evaluation of high-pressure hot-pressed TiB2and TiB2-20 vol pct Sic, conventionally hot-pressed Boride 2,' and KT silicon carbide. Boride

Jan 1, 1969

By W. W. Anderson

WEBSTER'S dictionary gives the following definition for "efficiency": "Effective operation as measured by a comparison of actual and possible results." Engineers think of this definition in terms of the ratio of output to input, and since they learn early in their careers that output cannot be greater than input, they usually consider a machine to be 100 pct efficient if it operates perfectly, and to be 0 pct efficient if it performs no function at all. Hence, values of efficiency between 0 and 100 pct reflect how well or how poorly a machine has performed; and any method of calculation for determining efficiency that does not show 0 pct when no function has been performed, and 100 pct when a perfect job has been done, is looked upon with skepticism. The literature on performance of coal cleaning equipment contains many references to efficiency of performance, but there is no unanimous agreement among the authors as to how the efficiency should be determined. Numerous methods of calculation have been proposed from time to time, most of which attempt to fulfill the definition of efficiency by comparing output to input, using float and sink data or some other characteristic, such as dry ash content, as a measure. However, to the best of our knowledge, none of the methods advocated in the past meets the requirements of a practical engineer in judging cleaning plant performance, because the formulas previously advocated do not confine numerical values of efficiency between 0 and 100 pct, and the literature does not contain an explanation of the meaning of the formulas under various conditions of operation. Therefore, the following formula, based on Yancey's concept of misplaced material,' is proposed as a measure of efficiency which will conform with the requirement that the values of efficiency must be confined between 0 and 100 pct: Quantitative Efficiency = 100.0 — (pct wt sink* in cleaned coal + pct wt float* in refuse)t • At gravity of separation effected by cleaning equipment. t Total product basis of the size of material under study. It will be noted from this formula that an actual separation between cleaned coal and refuse must be effected by the cleaning equipment before the formula can be used; also, that improperly distributed material in the products must be determined at the gravity of separation, rather than at some arbitrary test gravity. It will further be noted from this formula that the weight of cleaned coal recovered and the weight of refuse rejected by the equipment have their proper influence, because the sink in the cleaned coal and the float in the refuse are expressed in the equation as a per cent weight of the total products of a definite size discharged from the plant. This formula has one other important characteristic, in that it requires no knowledge of the feed to the cleaning unit. Comparison of data on the products from a cleaning unit with the feed to the unit can lead to many worthwhile conclusions. but not in regard to the efficiency of separation effected by the equipment. The reason for this is that mechanical fracture of many pieces of the feed occurs during the cleaning process; and in some processes crushers are deliberately incorporated in the flow to produce size degradation to free impurities from cleaned coal, thereby altering the washability characteristics of the material while in process. When such size degradation occurs in a cleaning process, it is quite possible that this crushing will result in a reduction in ash content in the finer sizes of cleaned coal when compared with the raw coal feed, even though no true separation has been effected. This is due to the fact that the finer sizes have been "sweetened" by the crushing of cleaned coal from some coarser size. Many formulas advocated in the past show high cleaning efficiencies

Jan 1, 1951

By P. L. Russell

The use of nuclear explosives for excavation has been demonstrated to be feasible and practical. Application of nuclear explosives for overburden removal from large ore deposits appears to offer economic advantages over present methods, but under the present pricing schedule, high-yield devices must be used to obtain an economic advantage. The use of nuclear explosives presents problems not inherent in the use of conventional explosives, and thus they can be used only in areas where large explosions and associated phenomena can be tolerated. Stripping of coal with nuclear explosives does not appear practical at this time, but stripping overburden from copper or iron ore deposits may be a potential use. Even here, however, the use of nuclear explosives would require isolation, favorable topography, favorable ground-water conditions, and a low population density. If these basic requirements can be met, the use of nuclear explosives for overburden removal would offer both economic and time advantages. The possibility of using nuclear explosives for peaceful purposes was a subject for speculation as early as 1945. The PLOWSHARE Program was formally established by the AEC in June 1957 for the purpose of developing and demonstrating peaceful applications for nuclear explosives. The technical direction of the program is the responsibility of the Lawrence Radiation Laboratory, Livermore, Calif. Mining and excavation were two fields that received early consideration as areas in which nuclear explosives might be used. Knowledge gained from underground weapons testing increased this possibility, and the development of the thermonuclear device which has a small fission-to-fission ratio, reduced the hazard of radiation, and spurred interest. Nuclear explosives appear to have the greatest potential in the following mining operations: 1.— Surface Mining a) Stripping of waste for open-pit excavation b) Breaking of ore in open-pit mining c) Quarrying d) Blasting for production of aggregate, riprap, etc. 2. — Underground Mining a) Block caving application b) Crushing of ore for in-situ leaching, etc. c) Preparation of underground waste disposal areas Of the potential applications, excavation is perhaps the most obvious and the best understood, and probably the most practical for use in the near future. Large quantities of rock and earth can be moved by proper emplacement of nuclear explosives, but economic and practical relationship must be determined between the known capabilities of nuclear explosives and the required excavating to remove overburden. REMOVING OVERBURDEN IN OPEN-PIT MINING Removing the overburden is one of the largest jobs in open-pit mining. In 1960 open-pit mines produced approximately 130 million tons of coal, 108 million tons of copper ore, and 135 million tons of iron ore. To obtain this much production, an estimated 3-billion tons of overburden were removed from the coal, 216 million tons from the copper ore, and 135 million tons from the iron ore. These figures do not include the initial stripping required before ore production. For example, the ASARCO Mission mine started stripping operations in August 1959 and by mid-November 1961 had removed 46.3 million tons of material including 1.2 million tons of ore.* Initial excavation of this type offers the most favorable conditions for utilizing the earth-moving ability of nuclear explosives. EXPLOSIVE CASTING FOR REMOVING OVERBURDEN The use of explosives to remove overburden is not a new idea. Conventional explosives have been used for this purpose for many years with varying degrees of success. Several small coal strip operators have used this idea to increase production. At these mines about one-third or more of the overburden is thrown by explosives across the pit and onto spoil piles. The additional costs of drilling more blast holes and more explosives are more than compensated for by a reduction in the time required to uncover the coal, and by an increase in coal production.

Jan 1, 1964

By W. C. Knopf, I. M. LeBaron

In the Carlsbad area potash is dry-mined and wet-concentrated. Wet concentration involves recircu-lation of saturated brines, with resultant difficulties of brine disposal and inherent losses in recovery through solution of the KCl. A dry beneficiation process for potash minerals would offer many advantages over present methods. International Minerals & Chemical Corp. has been working on development of a dry process for some time. This article is a progress report. Potash ores available in this country are largely mixtures of halite and sylvite contaminated with slimes and often with many sulfate and magnesium minerals. Electrostatic concentration of KCl when the contaminating minerals are in low percentage has been accomplished. Continuous production of better than 60 pct K2O concentrate from high potassium sulfate ores has not been demonstrated at the present time. Laboratory Procedure: Laboratory work on electrostatic beneficiation of potash minerals has been carried out using both pure crystals and mined ore. Electrostatic separation of the two main components of potash ore, sylvite (KCl) and halite (NaCl), is accomplished by first heat-treating the ore to allow for electric charging of the particles. Preferential exchange of' electrical charges between different mineral particles through surface contact is the principal mechanism involved. After proper conditioning of the particle surfaces, the oppositely charged minerals are dropped vertically into a horizontal electric field having a field gradient from about 5000 to 15,000 v per in. The particles are free-falling in the field. A suitable power supply is used to produce the proper total voltage for the electric field. Total required voltage is, of course, a function of electrode spacing. A typical electrode arrangement with power supply is shown in Fig. 1. In general, the total voltage on any system is 75,000 to 90,000 v. Power consumption is negligible, as leakage at insulators and corona discharge are the main losses in the system. Optimum preheat temperatures for pure material components generally are not the same as for mined minerals. Fig. 2 shows clearly that maximum recovery on a 50 pct pure KCl, 50 pct pure NaCl mixture occurs at a preheat temperature of about 600°F. In this test the preheat mixture was allowed to cool to about 225°F before dropping through the electric field. The optimum temperature for sep- aration is practically never the optimum temperature of preheat. Optimum separation for mined Carlsbad potash ores occurs when a preheat temperature closer to 900°F is used, as shown by Fig. 3, which plots percent sylvite (KCl) in the concentrate and tails as a function of preheat temperatures. Here again the ore was allowed to cool to about 225°F before separation. It should be noted that efficiency of separation increases continuously with increasing preheat temperatures until incipient fusion of the ore takes place at about 950°F. The increase in optimum preheat temperature for Carlsbad ores over the pure minerals can be traced to the effect of the clay slimes intimately associated with the mined potash. If the slimes associated with the ore are removed by washing with alcohol or other non-aqueous material, the necessity of preheating to the high temperature is removed. This temperature range has been proved critical within these limits during continuous pilot plant operation. Concentrate and tailings grade and optimum recovery also are functions of the temperature of the potash mineral particles during separation in the electric field. Fig. 4 presents data indicating that feed to the rougher pass should range between 150°F and 275°F for best results. It has been found by subsequent pilot plant work that these temperatures can be maintained in actual operation.

Jan 1, 1959

By C. L. Pillar

INTRODUCTION The location, size, character of the ore de- posit and its adjacent formations will deter- mine the mining system best used for its ex- traction. The mining system chosen will have to be engineered to suit the ore deposit and the people that will extract it. What can be done in the Philippines cannot always be accomplished in the upper Northwest Territory of Canada. Mining operations developed under a free enterprise system are planned to achieve a satisfactory return on the capital required to place them into production. Hence, the choices made in planning and developing the mining system must be dictated by sound economics. The mining system planned must favor: 1. Maximum ore extraction of the deposit with minimal grade dilution. 2. The amount of capital expenditure used to develop and equip the mine must be in harmony with the financial policy of the group yet be sufficient to provide a reliable mining system. Excessive frills and extras that are not required for efficient, safe production should not be provided. 3. Mine production costs are determined by the amount of men, material, equipment and energy used in the production process. These must be kept to minimal limits in order to insure dependable, low cost production. Extras that later are needed in the production operation can be added as required by the operator. Basically block caving is a mining system that is normally used to extract deep seated, massive, low grade deposits of copper, molybdenum and iron where stripping ratios for open cast mining are economically prohibited. The caving system is limited in its selectivity. It is a mass production type of underground ore extraction that under favorable conditions is one of the most efficient, low cost mining systems in the world today. Usually block cave mining systems are large. They are capable of producing ore at a rate of 10,000 to 100,000 tons/day from large deposits at costs that are comparable with many open cast mining systems burdened by high waste to ore ratios. Block cave mining systems differ basically in the plan of the area layout, the method of undercutting and the method of draw. The balance of the mining system is the mechanical transportation of the drawn ore to the surface, the mine supply, drainage and ventilation systems. This paper deals mainly with the methods of draw. The method of draw to be used in a block caving system dictates the development plan of the mining area, is highly influential in the method of undercutting to be employed and the supply and ventilation system to be used. There are three basic methods of draw commonly used in today's operations. They are: Gravity draw Slusher draw LHD draw The type of draw planned will have a major affect on: Capital expenditure and time required to bring a deposit into production. Productivity, efficiency and cost of the production operation. Percent recovery of the ore deposit in both tonnage and grade. Degree of production reliability Fundamentally the type-of draw used must be compatible with the physical characteristics of the deposit to be extracted and the financial position of the mining operation. The type of draw chosen must work to advantage with the size, configuration, depth, strength and fracture pattern of the deposit and its adjacent formations. It must favor the financial protection of the planned mining operation. A Gravity Method of Draw requires a finely fragmented, easy flowing product from the caving action of the deposit. To obtain the desired economic advantage the ore deposit must also have sufficient area and height of column to permit minimal development cost on a per ton basis. The gravity system re- quires the maximum amount of development openings and uses the minimum amount of mechanized equipment of all the caving systems. It is unskilled, labor-intensive and is highly sensitive to size of ore fragmen-

Jan 1, 1981

By P. A. Beck, Hsun Hu

It has been known for many years that in cold drawn polycrystalline aluminum the recrystallization texture is practically identical with the deformation texture.l,2,3 V. Goeler and Sachs4 stated that in cold rolled polycrystalline aluminum, too, the deformation texture is retained upon recrystallization. This behavior was interpreted by Dehlinger9 as an indication that each recrystallized grain inherits the orientation of the matrix in which it grew. However, Burgers and Basart showed6 that in deformed aluminum single crystals the orientation of the recrystallized grains strongly deviates from that of the matrix. Barrett demonstrated8 that even in deformed polycrystalline aggregates the local orientation within each deformed grain changes on recrystallization. No explanation has as yet been offered of how the aggregate texture can be retained, while the local texture is everywhere changing. According to Burgers and Louwerse6 the orientation change on recrystallization in compressed and rolled aluminum single crystals corresponds to a rotation around [112] axes lying in the active slip planes and perpendicular to active slip directions. This orientation change was rationalized by them on the assumption that the orientation of the recrystallized grains is determined by the orientation of the available nuclei, and that these nuclei are small fragments of the deformed crystal, rotated in the deformation process. Barrett pointed out,? however, that these assumptions were by no means proven, and that the total volume of material necessary for the nuclei is so small, that it must be available in any orien- tation whatever. It is, therefore, unlikely that an "oriented nucleation" theory could satisfactorily explain the orientation changes in recrystallization of a deformed crystal. Furthermore, in contradiction to Burgers and LOU-werse's conclusions, Barrett's results for the orientation change could be best expressed in terms of rotations around. [Ill] axes. Thus agreement is lacking not only in the interpretations but apparently even in the experimental results contributed by the various investigators. The orientation relationships on coarsening* in a highly oriented polycrystalline matrix were first investigated by Burgers and Basart.5 They produced a fine grained highly oriented matrix by the recrystallization of a compressed or rolled single crystal of aluminum, and studied the orientation relationships in three successive generations of crystals: the "deformation texture" in the single crystal, the recrystallization texture, and the coarsening texture. Burgers24 showed that the preferred orientation of each successive generation was almost always widely different from that of the preceding one, but that the coarse grains often approximately reverted to the orientation of their grandparent, the deformed single crystal. This observation, and the idea that the orientation of the coarse grains is determined by that of the available "nuclei," led to the concept that the coarse grains are nucleated by the remnants of the deformation texture.12,15 On the other hand, Bowles and Boas16 found, in a carefully conducted experiment, that in recrystallized fine grained copper of cube texture coarse grains grow in orientations which can be derived from the orientation of the recrystallized matrix by rotation around a [lll].] direction: the coarsening texture here does not appear to be related in any simple manner to the original rolling texture from which the cube texture had been obtained by recrystallization. It may be assumed that in a recrystallized material, even with a strong texture, there always are present some grains of practically any orientation. These may serve as "nuclei" for the coarse grains, provided that the conditions are favorable for their growth. The existence of a coarsening texture may be considered as an indication that grains with certain relative orientations with respect to the preferred orientation of their neighbors can grow much faster than others. This "oriented growth" theory was briefly suggested by van Arkel, as early as 1936,13 and it was mentioned by Burgers." Recently, C. G. Dunn produced direct

Jan 1, 1950

By J. C. Fisher

Nucleation theory is applied to martensite nucleation in substitutional iron alloys. Composition fluctuations are neglected, and a steady rate of nucleation is predicted for any composition and temperature. The maximum rate of nucleation (as a function of temperature) is shown to be measurable only for on extremely narrow composition range, being too great or too small outside this range. A number of experimental observations, including completely isothermal transformation in some alloys and composition-dependence of M. temperature in others, are compared with the calculated behavior. IN a previous report,¹ nucleation theory was applied to the formation of isothermal martensite in an Fe-Ni-Mn alloy. The experimental observations of Cech and Hollomon² were accounted for quantitatively. It now is of interest to examine the predictions of nucleation theory with respect to martensite transformation in substitutional iron alloys of other compositions. Since Fe-Ni alloys have received considerable attention, both as to their transformation kinetics and their free energies, they will be treated in detail, although the results apply with equal validity to other substitutional alloys. In substitutional alloys, where composition fluctuations can be neglected in volumes of critical nuclear size, the rate of nucleation of martensite in subcooled austenite is¹ n - Nv exp (—W*/kT) [I] where the free energy of formation of the critical size nucleus is W* = 8192 p (?² s³) [2] In the expressions for n and W*, the symbols have the following meanings: N is the number of atoms per unit volume; v, the atomic vibration frequency; 8, a function of the elastic constants of austenite and the shear angle of martensite; , the austenite/mar-tensite inter facial free energy; and ?fc is the free energy change per unit volume accompanying the transformation. Jones and Pumphrey² have determined the austenite to martensite free energy change as a function of nickel content and temperature for Fe-Ni alloys. Their expression is ?F = 2500 C + 1.25 (1 —C)?FFc cal per mol [3] where C is the atom fraction of nickel, and ?Ffe is the free energy change per unit volume for pure iron as tabulated by Zener4 (or, equivalently, Fisher"). From Eqs. 1 to 3, it is possible to calculate isothermal nucleation rates as a function of temperature and composition for Fe-Ni alloys. The only unknown parameter is (?² s³). The best value of this parameter for Fe-Ni alloys, determined from M. temperatures in a manner to be described shortly, is (?² s³)= 9.92 (10)21 cgs units [4] Taking this value of (?² s³), C-curves for isothermal nucleation were calculated for Fe-Ni alloys of several compositions, and are plotted in Fig. 1. It is evident from Fig. 1 that completely isothermal nucleation of martensite in Fe-Ni alloys can be observed experimentally for only a very narrow range of compositions. Cech and Hollomon find that a nucleation rate of about 10' nuclei per cc per sec is as fast as they can follow dilatometrically, and a rate of about 10-1 nucleus per cc per sec is about the slowest that can be measured conveniently. If the nucleation rate is to lie between the limits 10-1 10 the nickel content of an Fe-Ni alloy must fall in the atomic fraction range 0.299 5 C 5 0.302. Nickel contents lower than 0.299 correspond to nucleation rates too great to follow, and those higher than 0.302 to nucleation rates too slow to measure in a reasonable time.* In view of the narrow corn- position range in which successful isothermal measurements can be made, it is not surprising that only one substitutional alloy has been described' that comes close to being in the proper range.

Jan 1, 1954