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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 E. J. Koval

Practical miscible displacement processes will be characterized by lingering of the solvent into the oil. The lingering process is brought on by viscosity differences, and can be accentuated by channeling and longitudinal dispersion. The effects of these factors on the efficiency of unstable completely miscible displacements are combined in what is called the K-factor method. This method, analogous to the Buckley-Leverett method, predicts recovery and solvent cut as a function of pore volumes of solvent injected. Experimental data are included and show excellent agreement with theory for a wide variety of sandstone cores and viscosity ratios. INTRODUCTION Theoretical considerations,1 laboratory experiments,2 and pilot tests3 lead to the conclusion that miscible displacements in the field will be unstable In an unstable miscible displacement, the solvent fingers through the oil. This fingering leads to early breakthrough of the solvent and an extended period during which both oil and solvent are produced. For such a system, there appear to be four principal factors which bring about or accentuate the effects of instability: longitudinal dispersion (including geometrical effects), channeling, viscosity differences and gravity differences. Other factors, such as diffusion and flooding rate, can also influence the effects of instability, but at the flooding rates considered in this report, circa 15 ft/D, they are unimportant. Longitudinal dispersion can be thought of as a spreading of the solvent front caused by the presence of microscopic inhomogeneities. Channeling of the solvent occurs when a porous medium has macroscopic inhomogeneities; i.e., gross permeability variations. Viscosity differences lead to fingering of the less viscous solvent. This difference in viscosity accelerates the growth of fingers along paths previously developed because of permeability variations. Gravity differences lead to overriding by the usually less dense solvent. Al- though gravity effects are generally small at a flooding rate of 15 ft/D, they would, nevertheless, unnecessarily complicate the interpretation of flooding experiments. For this reason, all the experiments reported herein were done with matched density fluids. Fingering and the resultant poor areal sweep were recognized early as the dominant influences on the efficiency and the economics of miscible displacement processes. Much research effort has been spent on ways to minimize fingering and increase areal sweep such as the use of graded viscosity slugs4 or water slugs. Some researchers attempted to work out ways to prevent fingering completely; i.e., to achieve a stable displacement through gravity or latitudinal dispersion stabilization. l Others did not attempt to control fingering but obtained an economic process by merely recycling the solvent and sweeping pattern by pattern.6 During this period, all aspects of fingering came under close scrutiny. Some researchers reported on how fingering looks2 and how it is affected by viscosity ratio,7 geometry,7 and slug size.2,* Peaceman and Rachfordg suggested a mathematical approach to the prediction of unstable miscible displacements in relatively homogeneous sand packs, but their work cannot be extended conveniently to heterogeneous systems. Hence for a heterogeneous system, no method is presently available for predicting solvent cut and recovery as functions of pore volumes of solvent injected. The purpose of this investigation was to attempt to fill in the gap in our knowledge concerning the prediction of performance of unstable miscible dis placements. Necessarily, the system selected for study was a relatively simple one. The restrictions placed on the system were: (1) The system was linear; (2) The solvent was miscible in all proportions with the oil in place; (3) The solvent was continuously injected into the porous medium; (4) Gravitational effects were eliminated by matching densities and (5) All the flood rates were high and constant at 15 ft/D to avoid any small rate effect and to minimize any diffusion effects. To simplify and to indicate that both longitudinal dispersion and channeling arise from permeability variations, the effects which they cause or influence have been termed heterogeneity effects. The use of

By J. H. Scaff, H. C. Theurer, E. E. Schumacher

The microwave region of the radio spectrum was effectively utilized in radar designs during the recent war and it has become of increasing interest in the field of communications. Work in this field has led to an important use for silicon—that of the point contact rectifiers†1 in the frequency converter of microwave (radar or radio) receivers—and has stimulated considerable interest in the electrical properties and preparation of silicon and its alloys. Silicon is an electronic semiconductor. Its conductivity at room temperature results principally from the presence of certain impurities. While for metals an increase in impurity content increases the resistivity, for semiconductors such as silicon the opposite occurs and, in general, the addition of impurities lowers the resistivity. Silicon materials may be classified into one of two groups depending upon the manner in which the impurities contribute to electrical properties. These have been termed for convenience p-type or n-type. P-type silicon develops a very large positive thermal emf against metals, has a Hall coefficient of positive sign and a low resistance direction in point contact rectification with the silicon positive with respect to the point. N-type silicon, on the other hand, develops a negative thermal emf against metals, has a Hall coefficient of negative sign and a low resistance direction in rectification with the silicon negative with respect to the point. Impurities which produce n-type silicon are called donors inasmuch as these elements contribute to electrical conductivity by donating electrons to an unfilled energy band in the silicon. On the other hand, elements which produce p-type silicon are known as acceptors as these impurities contribute to the electrical conductivity by accepting electrons from a filled energy band permitting what is known as conductivity by "holes" in which the sign of the carriers appears to be positive. A general treatment of the mechanism of conduction in semiconductors from the viewpoint of modern band theory has been given recently by Pearson,2 by Becker, Green and Pearson,3 and by Torrey and Whitmer.4 In this investigation boron and aluminum have been found to be acceptors, and phosphorus, arsenic, and antimony to be donors in silicon. Data on the effect of boron and phosphorus on the electrical properties, when present singly and in combination, have been acquired. These data are discussed in the present paper. Raw Materials Used and Methods for Adding Seeond Constituents Silicon from two sources was used in this work and these will be referred to as A and B respectively. Silicon A is a material of high purity supplied by the Electro Metallurgical Co. It is prepared by chemical purification of "commercial" silicon obtained from the arc furnace reduction of SiO2. It contains 99.8 pct silicon, minimum, with small amounts of calcium, iron, aluminum, boron, and phosphorus as the principal impurities. This material was extensively employed in these studies as well as in the commercial preparation of rectifier materials. Silicon B is a material of high purity from the du-Pont Co. It is prepared by a pyrolytic reduction of Sicl4 and is free of analytically detectable amounts of boron and phosphorus. Its use permitted study of the effect of boron and phosphorus individually on the properties. This material contains, however, spec-troscopic traees of a number of metals, some of which affect electrical and rectification properties. Typical analyses for the two grades of silicon are given in Table 1. To make controlled additions of boron to the charge it was necessary to employ a low boron content master alloy since the quantity of boron to be added was usually only a few thousandths of one percent. The alloy containing nominally one percent boron was prepared by melting chemically pure boron with silicon A using the melting techniques for 320 g silicon

Jan 1, 1950

By Geo. A. Murphy

MR. ARTHUR E. GIBSON (General Manager, Consumers Mutual Coal Company of Utah) : I suppose the reason I was called upon to read this paper was because I was at the mine for several years, but I have not been there for the past four years and the installation has been put in since I left. This article has been written by my successor-it follows: At the very urgent solicitation of your Secretary I have consented to prepare this paper descriptive of the spiralizing plant which is a part of the new steel tipple recently erected by the Spring Canyon Coal Company to serve its mines at Spring Canyon, Utah. I have been reluctant in giving this consent for two reasons: one of which is an excellent one, and the other isn't any good at all. First,, I find that it takes a long time to become familiar with the eccentricities of spirals-to get wise to their whims and peculiarities, and while this plant has been in operation for a number of months, something new seems to crop up nearly every time I see it, and -it would be impossible to describe all of these things even if I remembered them. The second reason is, purely personal. Among the many things I have learned by attending the sessions of the Rocky Mountain Coal Mining Institute is that if a man is not pretty sure of the ground he attempts to cover, he had better keep his mouth shut, because if he isn't right, he is dead sure to be called down by someone. However, as I will not attend the session, this feature will not hurt much, but it will make it necessary for someone else to read the paper and this will make it much better for those who are obliged to listen to it because I don't read English very well. To begin with I will say that what is said in this paper applies only to the plant under discussion as I have had no experience with any other plant. While spirals have been in use for quite a number of years, and there are some large plants in the East, their use has not been general, and I believe the plant at Spring Canyon is the only one in the United States west of the Mississippi River. The details of the plant were worked out by Mr. Thos. R. Stockett, General Manager of the Spring Can= yon Coal Company; it was manufactured by the Anthracite Separation Company of Hazelton, Penn., incorporated in the design and erected by the Link-Belt Company of Chicago, in connection with the balance of the tipple. It consists of five units. Each unit consists of a vertical wood column about 6 inches in diameter, and 9 feet 6 inches in height, around which the spiral is built. It is something like a triple screw conveyor, set vertically, with flights down which the coal and refuse travel by gravity in the process of separation. There are three of these flights, forming a sort of nest but about 7 inches apart, and they make 21 turns around the column in its length. These flights dip toward the column about 27 degrees; they have a uniform pitch downward of 31 degrees; they are 19 inches wide at the top (measured from the column) and taper gradually until they are about 9 inches at the bottom. They are not made with one continuous plate from top to bottom, but are made up of a number of plates. Some of these plates are slightly waved, others have small corrugations in them and still others are perforated with small holes, all of which are intended to accelerate or. retard the movement of the coal and refuse as they pass down, At intervals there are adjustable plates in the flights for the purpose of decreasing the maximum outside;

Jan 1, 1926

By M. R. Geer, H. F. Yancey

John Grifien (Pittsburgh)—I wish to congratulate the authors on this paper, which, I am sure, will promote a clearer conception of the various criteria which have been advanced as measures of coal-cleaning performance. Their clear descriptions of the more important criteria that have been advanced and explanations of their meaning and application have been sorely needed and, if widely promulgated and followed, should remove much of the confusion which, heretofore, has been evident. I suggest that much of this confusion has also arisen from lack of common agreement as to the exact meaning of the words and terms used in our discussions of this subject. I am sure I have been equally guilty in this respect and suggest it is time that we select a committee to study vocabulary as well as criteria, so that common understanding and agreement can be had. Thus, I suggest, that the title of this paper should have used the words "coal-cleaning performance" and not "coal-washery performance." My understanding is that "washery" refers to a complete plant and it has additional functions to that of coal-cleaning, such as control of size, moisture, etc. Also, I do not agree with the authors that the formulas given in their paper, the Fraser and Yancey formulas, are necessarily the only ones that can be called efficiency formulas. When a dictionary is consulted one realizes that the words "efficiency" and "efficient" have not precise meanings but only very general ones and must be further defined for each particular use, if they are to have any precise meaning. To my mind, the first of the Fraser and Yancey formulas given should never be used as expressing "efficiency" alone but only by the fuller phrase which truly expresses the kind of efficiency it states, which is "coal recovery efficiency." For complete clarity we then need only define "coal" for the particular circumstances under which the formula is used. This would recognize the fact, fully explained by the authors, that the limits in coal qualities imposed by the market have imposed a factor—and a variable factor—which has influenced any answer obtained by this formula. Such a procedure makes it entirely practical and logical to use the values of any of the qualities of coal—ash, sulphur, phosphorus, specific gravity—or a combination of them to define what is coal for that particular use of the formula. Further, it is clear to me that the criteria to be used must be dictated by the objective in view and the use for which the data are obtained. The authors clearly point out that the Fraser and Yancey coal-recovery efficiency formula is influenced by three factors, "one is determined by the coal, one is dictated by market or use conditions, and one is an inherent characteristic of the cleaning unit itself." The formula which uses the inverse of misplaced material as a measure of efficiency is also a coal recovery efficiency formula and is influenced in like manner by these three factors. As the authors point out, both are useless in obtaining data which enables a direct and valid comparison of the performance of a given cleaning unit on two coals where the other two factors vary or of the performance of two or more cleaning units of different types unless the values of the other two factors are identical. Much of the confusion in our discussions of and literature on coal cleaning has come from failure to understand clearly this fact. Many such comparisons have been made from which erroneous conclusions have been drawn. I am, therefore, particularly glad to see the authors' emphasis on 4?-1e criteria which indicate the "sharpness of separation. I feel that these are the essential criteria

Jan 1, 1952

By C. W. Sherman, John Chipman

IN the mathematical statement of the law of mass action, the activity of each substance consumed or produced in a reaction is used to obtain a numerical constant which is characteristic of the equilibrium of that reaction at the given temperature. The activity in many cases is equal to, or proportional to, the concentration. If this is the case, any units denoting concentrations, such as weight percentage or mol fraction, may be substituted in the equilibrium expression. It is even permissible to use mixed units provided only that each unit is that used in defining the activity of the corresponding substance. The indiscriminate use of concentration units in writing equilibrium constants sometimes leads to very erroneous results. On the other hand, what appear to be contradictory results from similar investigations often might be resolved if the activity is the quantity that is used to determine the equilibrium constant for the reaction in question. De-sulphurization serves as an important illustration. Sulphur in lron and Steel The control of the sulphur content of the finished product is one of the oldest and broadest problems in the metallurgy of iron and steel. In general, this control must be exercised through the action of a slag, and for this reason experimental studies of the distribution of sulphur between slag and metal have assumed considerable importance. Under slags of the type found in the basic open-hearth furnace the ratio (% S)/[% S] does not in general exceed about 8. Under blast-furnace conditions, however, de-sulphurization ratios of 50 to 100 are readily obtained, and under laboratory conditions using similar slags, the ratio may be as high as 400. The fact that blast-furnace metal is much more readily desulphurized than is molten steel rests upon two distinct phenomena. The first of these is connected with the slag reaction. In the open-hearth furnace, the equilibrium in the following reaction places a definite limit on the sulphur content of the slag: S + 0-- = 0 + S- - where the symbols underlined indicate elements dissolved in the liquid metal, while slag constituents are represented as ions. The high oxygen content of the open-hearth bath prevents this reaction from proceeding far enough to effect good desul-phurization. In the blast-furnace process, however, the oxygen activity is limited by the presence of carbon in the metal and in the coke bed and the controlling reaction becomes: S_+O- + C_=S- + CO(gas) [2] The second phenomenon which contributes to the easier desulphurization of blast-furnace metal is the effect of high concentrations of other elements. These elements, notably carbon, silicon, and phosphorus, increase the thermodynamic activity of sulphur several fold. As a result of this increase, the escaping tendency of sulphur from the metal phase is increased and its removal facilitated. Previous Studies It becomes, therefore, a matter of importance to determine the effects of various elements upon the activity of sulphur in the melt. The experimental approach to this problem has been made through the equilibrium between sulphur in the melt and gaseous atmospheres containing hydrogen and hydrogen sulphide. The reaction and its equilibrium constant are written as follows: S + H2(gas) =H2S(gas);K1 = -- [3] Pn2 • as where the symbol a, represents the activity of the sulphur in solution in the metal. The understanding of the chemical behavior of sulphur in steel and pig iron rests on a firm knowledge of the binary system Fe-S. A number of studies of the equilibrium have given closely agreeing results from which the activity and free energy of the dissolved sulphur may be found. The recent paper by Sherman, Elvander, and Chipman1 includes a discussion and quantitative comparison of the results of several of the significant researchesr-" on dilute solutions of sulphur in iron. The results of this study indicate that the relationship between the ratio Ph,s/p+ and the percentage of sulphur in solution is not a linear one. This

Jan 1, 1953

By F. Gordon Smith

ALTHOUGH several geological indicators of the critical type are known, including quartz inversions and decomposition of hydrous minerals such as serpentine, there are very few of the general type. Solid solutions are excepted, but the limitations of use are very restricted and interpretations are sometimes ambiguous. General methods for determining temperature and pressure conditions during the crystallization of minerals would have considerable scientific and economic value. It is not the purpose of this paper .to, discuss the various methods of geological thermometry and barometry, but to present one general method, applicable to all minerals, and to describe what progress has been made in the methods of measurement. The general method, in. brief, is a study of the stress conditions in and around various types of foreign inclusions which are trapped in minerals during growth. The method depends upon the fact that any homogeneous gas, liquid, or solid will in general have coefficients of thermal expansion and volume compressibility different from those of any given mineral. Therefore stress must develop in and around all types of inclusions in minerals if the temperature or the pressure, or both, are changed from the conditions which prevailed during the growth of the mineral. The methods of measurement consist of determinations of the temperature-pressure conditions of fit of the inclusions in the host mineral. The types of inclusions in minerals are: 1-gas, or liquid plus vapor, when observed at room temperature, due to crystallization under pneumatolytic conditions; 2-liquid, or liquid plus vapor when observed at room temperature, due to crystallization under hydrothermal conditions; 3-glassy solid, or devitrified glass, due to crystallization under magmatic or high temperature metamorphic conditions, in a siliceous liquid; and 4-crystals, due to overgrowth of other minerals crystallizing simultaneously or of other minerals which crystallized previously. A survey of the literature shows that much valuable earlier work on inclusions, especially that carried out in England in the last century, has dropped out of current knowledge. The following is a brief summary of the significant, contributions to the problem up to the present day. Davy in 1822 asserted that fluid inclusions in minerals consist of an aqueous solution of salts and a gas bubble, the whole being either at lower or higher pressure than atmospheric.1 At intervals from 1823 to 1862, Brewster contributed information concerning other types of inclusion consisting of 1-aqueous solution, a much more expansible liquid, and a gas; 2-aqueous solution, salt crystals, and a gas; and 3-the very expansible liquid and a gas. The very expansible liquid fills the gas space between 20° and 30°C. Compression strain effects were seen around inclusions in diamond, topaz, and other minerals.2-6 Sorby in 1858 and 1869 further advanced the study begun by Davy, stating that fluid inclusions represent a sample of the mother liquor of crystallization and that the degree of filling of aqueous inclusions at room temperature defines the temperature-pressure relations during formation of the host. The degree of filling may be measured by determining the minimum temperature of filling of the inclusion by the liquid phase. The very expansible liquid in some fluid inclusions is liquid. carbon dioxide. The temperature at which salt crystals in fluid inclusions completely dissolve in the fluid is the minimum temperature of formation. Inclusions of glass or devitrified glass indicate crystallization from a melt. Inclusions of crystals in minerals are often centers of strain, which may be seen by optical effects or by radial tension cracks. Sorby realized that an analysis of stress-strain relations about inclusions could be used to provide precise data on the temperature of crystallization, but the matter was never pursued.7,8 Hartley (1876, 1877),9,10 Hawes (1881),11 Wright (1881),12 Johnsen (1920),13 and Holden (1925)14

Jan 1, 1952

By E. L. Dougherty

A phenomenological theory for a one-dimensional unstable miscible displacement similar in type to the Buckley-Leverett model but including the effects of mixing is proposed An equation giving the fractional flow of pure solvent through an oil phase containing dissolved solvent is derived including the effects of gravity and nonbomogeneities. Numer ical integration of the resulting pair of nonlinear hyperbolic partial differential equations gives volumes of dissolved and undissolved solvent as functions of space and time. Parameters in the model which characterize mixing due to dispersion were determined from experimental data; the parmeters correlated with viscosity ratio. The results indicate that the rate of dispersive mixing is proportional to volumetric flow rate Incorporated into our equations were Koval's heterogeneity factor H. which characterizes mixing due to channeling. This allowed satisfactory predictions of displacement behavior observed in horizontal floods in nonhomo-geneous cores; but in most cases the values of H which we used were considerably larger than those used by Koval. INTRODUCTION It has been shownl,2,3 that for favorable viscosity ratios, the diffusion equation with convection satisfactorily describes the behavior of a miscible displacement in a porous media. Numerous attempts to develop a satisfactory mathematical description for the case of unfavorable viscosity ratio have been reported, but they have met with only partial success. These attempts, which are reviewed by oval! have taken two approaches: (1) development of a flow model akin to the Buckley-Leverett approach for immiscible displacement neglecting the effects of mixing5,6 and (2) simultaneous application of Darcy's Law for flow and the diffusion equation with a convection term for mass transfer.7,8,9 We set out to construct a mathematical model of Type 1 including, though, the effects of mixing. Based upon a set of hypotheses, we derived a pair of nonlinear hyperbolic partial differential equations which describe in a one-dimensional system the combined effects of flow and mixing. To work with these equations it was necessary to develop a fractional flow formula. The combined system was integrated numerically using the method of characteristics. In our equations the mixing process is characterized by four parameters. Three of these, labeled ß, P1 and P2, account for dispersive type mixing. The fourth is the heterogeneity factor H, proposed by Koval to account for channeling due to nonhomogeneities in the porous media. Values of the parameters which would cause agreement between theory and experiment were determined by trial-and-error for miscible floods conducted in horizontal cores of both homogeneous and heterogeneous materials. Calculations were also performed for vertical floods in homogeneous cores. The purpose of this paper is (1) to present the details of the mathematical analysis, (2) to present the results of the calculations, (3) to consider what light the results shed on the mixing process in an unstable miscible displacement, and (4) to provide a firmer foundation for the correlation technique developed by Koval for predicting the behavior of unstable miscible floods. STATEMENT OF PROBLEM The problem is to construct a mathematical model which describes in one dimension the observed behavior in an unstable miscible displacement. The approach is phenomenological in that the equations are based on assumptions which violate certain physical precepts known to apply to the displacement phenomenon. However, the assumptions do allow us to account mathematically for the more essential phenomena. The work is of value if we can quantitatively predict experimental results which heretofore could not be predicted, even though the synthetic nature of the model belies complete explanation of the observations. We assume that the system is comprised of two contiguous flowing phases. One phase, which we call free solvent, has the properties of pure solvent. We call the fraction of the pore volume occupied by this phase the solvent saturation, designated s. The oil phase consists of oil containing dissolved

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. B. Mee

The secondary re crystallization behavior of two vacuum-melted, nominally 3 pct Si-Fe alloys has been studied, one alloy being produced from pure electrolytic iron and semiconductor-grade silicon and the other from commercial-purity hot band. The vacuum-melted and vacuum-remelted alloys were found to develop {110}(hhl) and {lll)(hkl), respectively, on annealing in high vacuum and {100}(0kl) and {11l}(hkl), respectively on annealing in pure hydrogen of dew point, < — 60°C. H2S doping of the hydrogen or annealing in commercial argon was found to effect a reversal for the vacuum-remelted alloy from (111) to (100) secondary recrys-tallization, while prolonged vacuum annealing at 2 x 10-6 Tow resulted in a tertiary re crystallization from (111} to (110). The growth selection of the {lll}(hkl) component was found to be governed predominantly by the {111}/planar misorientation, the maximum deviation of the resulting secondaries being 5 deg. It is considered that the driving force for {111}(hkl) secondary recrystallization arises from the preferential lowering of ?111, possibly by phosphorus or nitrogen diffusion, just as oxygen and sulfur under certain conditions have been found to preferentially lower ?100. When heavily cold-rolled metals are annealed at high temperatures, primary recrystallization usually occurs rapidly and may be followed by considerable grain growth. However, in the presence of a fine dispersion of inclusions1"3 or with a strong matrix preferred orientation4&apos;&apos; the matrix may be stabilized and a secondary recrystallization may then be effected. Also, in vacuum-melted Si-Fe it has been shown possible to develop {100}(001) and {110}(001)6 surface energy driven secondary recrystallization textures. Further, it has been found that, on annealing very pure metals, grain growth normally ceases when a grain size approximately twice the strip thickness is acquired, due to the formation of V-shaped thermal grooves which result in an equilibrium being set up between surface and grain boundary stresses. If, however, a surface energy difference ??s, exists between adjacent grains (of the order of a few percent of the grain boundary energy ?b), then according to Mullins8 a secondary recrystallization may be effected, since the grains having lowest ?s would grow at the expense of their higher-energy neighbors. For clean bee metal surfaces ?110 should be lowest; however, for 3 pct Fe-Si (100) secondary structures have frequently been developed and have been attributed by Walter and Dunn9 to an alteration of the relative surface energies due to the preferential chemisorption of oxygen on (100) planes. Kohler,10 however, indicated that this reversal may be achieved by employing minute quantities of highly polar compounds in the annealing atmosphere, i.e., sulfur compounds, in par- ticular H2S being beneficial in (100) growth. This work has been verified by Kramer" using {100)/{110} bicrystals, while Mager12 demonstrated the importance of sulfur diffusion within the strip. Detert13 has intimated that carbon may be responsible for the preferential lowering of ?110 although no direct evidence has appeared yet to support this supposition. It must , be appreciated that the above-mentioned examples of surface energy driven secondary recrystallization were obtained with very pure vacuum-melted materials with oxygen contents (which is now known to be a critical factor") less than 35 ppm. During the course of a textural study on Si-Fe it has been found that under certain conditions (111) secondary recrystallization can be obtained by what is believed to be a surface energy growth selection mechanism. The purpose of the present paper is to briefly describe this result and to try and postulate the possible causes. EXPERIMENTAL Two vacuum-induction-melted, nominally 3 pct Si-Fe alloys were produced, one from pure electrolytic iron and semiconductor-grade silicon designated A and the other from commercial-purity hot band, designated B.

Jan 1, 1969

By Walter H. Munk

Generation of Waves by storms, and the transformation of waves in shallow water by local bottom topography are briefly reviewed. A detailed description of water motion in waves explains the nature and distribution of wave forces. A numerical example dealing with forces and moments against vertical piles summarizes the principles involved. Introduction The oil industry is moving into the open waters of the Gulf of Mexico. The design and construction of offshore drilling rigs present a number of difficult problems, among them problems involving the action of waves on fixed and floating structures. This paper deals with our present knowledge of ocean waves as it pertains to these problems, and the principles involved in computing wave forces against fixed structures. It is hoped that the more difficult case pertaining to floating structures can be discussed at a later time. None of the engincering aspects of the problem are entered upon in this paper, yet an understanding of the principles presented here should be helpful in the solution of practical problems. Waves are capable of exerting almost unbelievable forces. During a gale, waves at the harbor entrance of Amsterdam Canal in Holland lifted a 20-ton concrete block, which was resting on the bottom in 12ft of water, and deposited it on top of a pier which was 5 ft above the high water mark. At Wick Breakwater in Scotland, a mass of large stones in cement and bound together with iron rods, weighing 1350 tons, was broken loose and moved bodily! But one need not go to Holland or Scotland to find evidence of damage that waves are capable of inflicting (Fig I and 2). What is the mechanism by which waves are capable of exerting such enormous forces? To answer this, we must consider briefly the nature of wave motion. When a wind blows over a wheat field, waves appear to travel across the field, whereas the wheat itself remains rooted, slowly swinging back and forth. Similarly, in the case of water waves, the waveform travels swiftly over the water surface, whereas the water particles oscillate back and forth, but scarcely advance. The speed of the wave form is called wave velocity, the speed of the water particles is called orbital velocity. This fundamental difference between wave velocity and orbital velocity was already recognized by Leonardo da Vinci. Orbital velocity is generally much slower than wave velocity except for a breaking wave, a case which is of particular importance in the application to engineering problems. Forces exerted by the waves on obstacles

Jan 1, 1949

By J. P. Matoney, E. J. Sandy, D. A. Dahlstorm

INTRODUCTION Before an understandable discussion concerning coal-dewatering processes can be initiated, it is necessary to define some terms of usage. The many different meanings and interpretations of terms preclude development of standard definitions so that the terms which are defined in this chapter should be regarded mainly as working terms intended to facilitate communication. A coal slurry &apos; might be defined as a watery suspension carrying particles of fine coal and other assorted material ranging upward in size from the finest clay particles. Some particles may be in the colloidal sue range and exhibit characteristics of suspended colloids. The dilution of the system generally is such that it is in a condition of independent particle subsidence. In some coal mining areas a slurry is regarded as a suspension of fine coal and other materials having a dilution that cannot be economically dewatered by filtration. A coal slime may be regarded as a slurry containing particles of such size range that 50 per cent or more (by weight) will pass a 200-mesh sieve. Slime suspensions generally are difficult to handle without special treatment in conventional sedimentation equipment. A coal sludge is defined as a slurry that has been partly dewatered by sedimentation usually to a thick or viscous consistency where further dewatering can be most readily facilitated by filtration. Culm and silt are terms characteristic of coal-preparation practice which should be defined. In anthracite terminology, culm is the generally coarse residue or waste accumulation of coal, bone and rock from preparation plants most of which are no longer functional. Culm has a wide range of sizes, some rather coarse, and except to differentiate it from silt, the term is of limited interest in this discussion. In anthracite terminology, silt is the accumulation of waste fine coal, bone and slate settled out from preparation plant water in excavated basins. Silt is generally made up of particles ranging in size from 3/32-in. round-opening screen to the very finest sues. Silt is also referred to at times as sludge, fines, slush, and mud and is the partly dewatered derivative of what has been previously defined as slurry. In the terminology of the bituminous coal producing areas, the word culm commonly corresponds to slurry or slime, depending upon the size distribution of the suspended solids. Editor&apos;s Note: Because of the numerous and varied contributions of the associate authors of this chapter, the text does not differentiate between sections authored by each individual as is the case in most chapters of this book.

Jan 1, 1968

By W. R. McMillan, E. A. Gulbransen, K. F. Andrew

Annealed and electrolytically polished pure iron was oxidized between 650° and 850°C at oxygen pressures of 0.1 to 2 microns Hg. Electron optical studies showed that oxidation occurs discontinuously over the surface and orients crystallographically with the substructure of the metal grains. IN previous papers&apos;, " the authors described the formation of oxide films on high purity iron at norrnal pressures and at temperatures below 300°C. Electron optical studies showed that the oxide film consisted of many individual oxide crystallites having nearly random orientation on each metal grain. The size of the oxide crystallites depended upon the temperature and upon the extent of oxidation. In these studies the crystallite size varied from 250 to 1800Å. Recently, Bardolle and Benard3 studied the crystal habit of the oxide crystallites formed on Armco iron between 650" and 850°C in what might be defined as a "poor vacuum" atmosphere in which the pressure could be varied between 10-1 to 10-3 mm of Hg. Their results showed that, at 850°C and a vacuum of l0-2 to 10-3 mm of Hg, a few well oriented nuclei of oxide were formed on the metal grains. At pressures between l0-1 to 10-2 mm of Hg, many oriented nuclei of oxide were formed. At 750°C the metal surface was bright although an oxide was present. Unfortunately, the extent of oxidation, the oxygen pressure, and the influence of carbon in the metal were not determined, and it was, therefore, impossible to give a theoretical interpretation to the very beautiful light micrographs. Bardolle and Benard used the light microscope to determine the crystal habit of the oxide and X-ray diffraction to determine the orientation relationships. It was believed that the use of controlled amounts of oxidation at high temperature is an important method for the study of the initial stages of oxidation of metals at high temperatures. This paper presents the first results of an electron optical study of this problem for iron. Since the electron microscope and electron-diffraction methods using stripped oxide films and surface replicas yielded finer details of the crystal habit and structure of oxide film, these methods were used in this study. Two types of iron having different carbon contents were used: Puron and Armco* iron. A com- parison of the results for these materials made possible a study of the influence of the surface oxide-carbon reaction on the initial stages of oxidation. Before discussing the experimental results, the thermodynamic equilibria of 1—the oxidation reaction, 2—the surface oxide-carbon reaction, and 3—the solubility of oxygen in iron, will be considered. In addition, it is important to consider the kinetic-theory predictions on the rate of collision of oxygen with the iron surface. Thermodynamic and Kinetic-Theory Calculations Only one type of reaction is usually considered in the oxidation of iron at normal pressures at high temperatures: Fe(a) + 1/2O2(g) = FeO(s) 3Fe(a) + 2O2(g) = Fe3O4(s) 2Fe(a) +3/2O2(g) = a-Fe2O3(s). [11 However, two other reactions may occur: FeO(s) + C (solid solution in Fe) + Fe(a) + CO(g) [2] and Fe(a) + O2(g) e Fe(a) (solid solution of 0,). [ 3 ] In addition, iron will react with the higher oxide at certain temperatures to form FeO. For a complete interpretation of the oxidation of iron at low pres-

Jan 1, 1955

By Richard R. Wilding

The United States Coal Mining Indus- try had the opportunity to greatly improve their roof control techniques during the nineteen seventies by the use of fully grouted resin roof bolts. This new tool was first approved for underground coal mines by M.S.H.A. ten years ago and during the next decade has been perfected to meet many problems previously thought to be unsolveable. The initial approved cartridge was very similar to the ones used in Europe in the nineteen sixties and since then has undergone a great deal of modification for our mines. These refinements in cartridge construction have been obtained by the close cooperation of resin manufacturers with mining personnel in perfecting pro- ducts customized to particular needs. This type of continuing program naturally necessitates a responsibility upon the resin anchor manufacturer to maintain a competant technical department. These roof control technicians by constant exposure to many of the most demanding underground problems, have developed through the years into leaders in successful applications. Today, Celtite has over twenty technically trained underground technicians with an accumulation of over one hundred years experience in resin anchors. Our documented technical reports are in excess of five thousand, involving over six hundred mines representing every major coal mining area in most mineable seams. This department has been the implement used to work with our research and development people in perfecting a continual line of improved resin anchors. This program has been a significant factor in leading the resin anchor industry into a position where approximately one third of the 1980 under- ground coal production used the fully grouted system. However, as gratifying as this accomplishment is, our experience has lead us to believe that there is an additional demand for a roof control system capable of combining the mechanical bolts ability to apply tension with resins proven superiority in anchoring. A significant number mines visited during the last decade had problems that we believed could be better answered by applying tension rather than the fully grouted approach. In most in- stances, their current mechanical installations were doing an adequate job until anchor slippage occured. Fully grouted systems solved this bleed-off problem. However, an acceptable number of falls still occured. In these some- what unique areas which we believe may amount to as much as fifteen percent of the total bolt installations, the need for improved anchors was a necessity. In developing this program we listed the ultimate requirements necessary for both the bolt and the anchor. These are for the bolt: A. Have one end manufactured so it can be grouted to the mine roof with polyester resin, obtaining a variable length from 12" to 36". B. Have a mechanism which allows the operator to tension the roof from zero to 10,000 psi. This mechanism should indicate if the roof is in compression and the bolt in tension. C. Have a mechanism which allows the operator to thoroughly mix the resin. D. The head of the bolt should not extend below the roof line any more than a normal mechanical bolt. E. Not require any extra tools for installation. F. Be bendable for use in low seam coal. G. Be durable and inexpensive. The ultimate point anchor resin should have the following characteristics: A. Have a mix and gel cycle which would compete with a mechanical bolt for speed of installation. B. Have a pull out strength in excess of a normal 5/8" mechanical bolt.

Jan 1, 1981

By L. R. Page, J. J. Norton

EFFECTIVE methods for determining grade and reserves of pegmatites in advance of mining have been developed in recent years. When intensive work began on the economic geology of pegmatites during the early years of World War 11, it was widely believed that grade and reserves could be estimated only in a most general way. The literature indicated that not all pegmatite deposits are unpredictable, but statements emphasizing unpredictability and irregularity seem to have been more widely accepted. Many standard references reflect the prevailing opinion that minerals: are unsystematic in their occurrence in pegmatites. Pegmatites have been described as "notorious for their irregular structure," and minerals within pegmatites as "irregularly distributed." "irregularly disseminated," of "erratic occurrence," "proverbially erratic in distribution," "characterized by extreme irregularity in ... size and distribution of the essential minerals, and variety and proportion of the accessory minerals," and "spotty.&apos;"" Over many years, however, geologic techniques have been applied with increasing success to economic analysis of pegmatites. Investigations, chiefly since 1939, have shown that mineral distribution and structure of pegmatites are largely systematic and that detailed geologic study can be used to predict the size and shape of mineral deposits in pegmatites. Because the writers are most familiar with the pegmatites of the Black Hills, S. D., examples cited are from that area." Pegmatites mined in other dis- * All examples are described more cornpletely elsewhere.: tricts are similar to those in the Black Hills," and the techniques described here have been applied by geologists and engineers working in other districts. In actual practice few pegmatite mine operators attempt to make accurate estimates of grade and reserves in advance of mining. Many deposits are too small for mining except on a very small scale; few pegmatite mines have more than ten employes. Hence there is little systematic preparation for mining, and the capital outlay may be so small that only a general knowledge of reserves is necessary. Technical personnel are rarely employed by pegmatite mine operators to make such estimates. In the future, however, as capital outlay becomes larger, estimates of grade and reserves will have to be more accurate. Geologic Aspects of Determining Grade and Reserves Estimating grade and reserves requires that the type of pegmatite, the size and shape, and the internal structure be determined by detailed geologic mapping and careful mineralogic study. The wide variations in the texture and distribution of minerals within many pegmatite deposits make a thorough geologic study of this kind necessary before it is possible to decide what sampling methods should be used to determine mineral content or grade. Types of Pegmatites: The first step in economic appraisal of a pegmatite is to determine the min-eralogical and structural type. In general, pegmatites may be classified into two groups: 1) those in which component minerals are more or less evenly distributed throughout the entire body and show relatively minor variations in texture, and 2) those in which component minerals have been segregated into units with recognizable differences in texture, mineralogy. or both. Johnston" has classified these types of pegmatites as homogeneous and heterogeneous, respectively. Cameron et al.10 have used the terms ~inzoned and zoned for similar types. The difference between the homogeneous and heterogeneous types is of great commercial importance. Homogeneous pegmatites can seldom be mined where the products are concentrated by the hand methods now commonly used, but they have been mined in a few places where a milling process has

Jan 1, 1957

By John J. Sammarco

This report (Safety File Guidance 6.0) is the eighth in a nine-part series of recommendations and guidance addressing the functional safety of processor-controlled mining equipment. It is part of a risk-based system safety process encompassing hardware, software, humans, and the operating environment for the equipment’s life cycle. Figure 1 shows a safety framework containing these recommendations. The reports in this series address the various life cycle stages of inception, design, approval and certification, commissioning, operation, maintenance, and decommissioning. These recommendations were developed as a joint project between the National Institute for Occupational Safety and Health and the Mine Safety and Health Administration. They are intended for use by mining companies, original equipment manufacturers, and aftermarket suppliers to these mining companies. Users of these reports are expected to consider the set in total during the design cycle. • 1.0 Safety Introduction (Part 1).—This is an introductory report for the general mining industry. It provides basic system/software safety concepts, discusses the need for mining to address the functional safety of programmable electronics (PE), and includes the benefits of implementing a system/software safety program. • 2.1 System Safety (Part 2) and 2.2 Software Safety (Part 3).—These reports draw heavily from International Electrotechnical Commission (IEC) standard IEC 61508 [IEC 1998a,b,c,d,e,f,g]and other standards. The scope is “surface and underground safety-related mining systems employing embedded, networked, and nonnetworked programmable electronics.” System safety seeks to design safety into all phases of the entire system. Software is a subsystem; thus, software safety is a part of the system’s safety. • 3.0 Safety File (Part 4).—This report contains the documentation that demonstrates the level of safety built into the system and identifies limitations for the system’s use and operation. In essence, it is a “proof of safety” that the system and its operation meet the appropriate level of safety for the intended application. It starts from the beginning of the design, is maintained during the full life cycle of the system, and provides administrative support for the safety program of the full system. • 4.0 Safety Assessment (Part 5).—The independent assessment of the safety file is addressed. It establishes consistent methods to determine the completeness and suitability of safety evidence and justifications. This assessment could be conducted by an independent third party. • Safety Framework Guidance.—It is intended to supplement the safety framework reports with guidance providing users with additional information. The purpose is to assist users in applying the concepts presented. In other words, the safety framework is what needs to be done and the guidance is how it can be done. The guidance information reinforces the concepts, describes various methodologies that can be used, and gives examples and references. It also gives information on the benefits and drawbacks of various methodologies. The guidance reports are not intended to promote a single methodology or to be an exhaustive treatment of the subject material. They provide information and references so that the user can more intelligently choose and implement the appropriate methodologies given the user’s application and capabilities. The guidance reports comprise parts 6 through 9 of the series and are listed below: [5.1System Safety Guidance (Part 6).—This guidance supplements 2.1 System Safety. < 5.2 Software Safety Guidance (Part 7).—This guidance supplements 2.2 Software Safety. < 6.0 Safety File Guidance (Part 8).—This guidance supplements 3.0 Safety File. < 7.0 Independent Functional Safety Assessment Guidance (Part 9).—This guidance supplements 4.0 Independent Functional Safety Assessment.] [ ]

Jan 4, 2006

By John J. Sammarco

This report (System Safety Guidance 5.1) is the sixth in a nine-part series of recommendations and guidance addressing the functional safety of processor-controlled mining equipment. It is part of a risk-based system safety process encompassing hardware, software, humans, and the operating environment for the equipment’s life cycle. Figure 1 shows a safety framework containing these recommendations. The reports in this series address the various lifecycle stages of inception, design, approval and certification, commissioning, operation, maintenance, and decommissioning. These recommendations were developed as a joint project between the National Institute for Occupational Safety and Health and the Mine Safety and Health Administration. They are intended for use by mining companies, original equipment manufacturers, and aftermarket suppliers to these mining companies. Users of these reports are expected to consider the set in total during the design cycle. • 1.0 Safety Introduction (Part 1).—This is an introductory report for the general mining industry. It provides basic system/software safety concepts, discusses the need for mining to address the functional safety of programmable electronics (PE), and includes the benefits of implementing system/software safety program. • 2.1 System Safety (Part 2) and 2.2 Software Safety (Part 3).—These reports draw heavily from International Electrotechnical Commission (IEC) standard IEC 61508 [IEC 1998a,b,c,d,e,f,g]and other standards. The scope is “surface and underground safety-related mining systems employing embedded, networked, and nonnetworked programmable electronics.” System safety seeks to design safety into all phases of the entire system. Software is a subsystem; thus, software safety is a part of the system’s safety. • 3.0 Safety File (Part 4).—This report contains the documentation that demonstrates the level of safety built into the system and identifies limitations for the system’s use and operation. In essence, it is a “proof of safety” that the system and its operation meet the appropriate level of safety for the intended application. It starts from the beginning of the design, is maintained during the full life cycle of the system, and provides administrative support for the safety program of the full system. • 4.0 Safety Assessment (Part 5).—The independent assessment of the safety file is addressed. It establishes consistent methods to determine the completeness and suitability of safety evidence and justifications. This assessment could be conducted by an independent third party. • Safety Framework Guidance.—It is intended to supplement the safety framework reports with guidance providing users with additional information. The purpose is to assist users in applying the concepts presented. In other words, the safety framework is what needs to be done and the guidance is how it can be done. The guidance information reinforces the concepts, describes various methodologies that can be used, and gives examples and references. It also gives information on the benefits and drawbacks of various methodologies. The guidance reports are not intended to promote a single methodology or to be an exhaustive treatment of the subject material. They provide information and references so that the user can more intelligently choose and implement the appropriate methodologies given the user’s application and capabilities. The guidance reports comprise parts 6 through 9 of the series and are listed below: [< 5.1 System Safety Guidance (Part 6).—This guidance supplements 2.1 System Safety. < 5.2 Software Safety Guidance (Part 7).—This guidance supplements 2.2 Software Safety. < 6.0 Safety File Guidance (Part 8).—This guidance supplements 3.0 Safety File. < 7.0 Independent Functional Safety Assessment Guidance (Part 9).—This guidance supplements 4.0 Independent Functional Safety Assessment.] [ ]

Jan 8, 2005

By F. I. Stalkup

Vapor-liquid phase equilibrium experiments have been conducted in a static equilibrium cell on mixtures of a light, 45 API stock- tank gravity reservoir fluid and a rich hydrocarbon gas containing approximately 55 mole per cent of intermediate hydrocarbons. Both apressure-vs-composition study of the gas and a simulated reservoir fluid, and a multiple-batch contact simulation of the condensing-gas-drive oil recovery process were performed. In the latter experiments equilibrium gas and liquid compositions were analyzed. Also, conventional, "condensing-gas-drive", long-tube displacement experiments of the reservoir fluid and gases of various richness were performed. The results of these experiments could not be satisfactorily interpreted by the conventional pseudo-ternary-diagram representation of multicom-ponent phase behavior. The results seem to be explained better by considering a bubble-point surface and a dew - point surface joined in a plait-point locus. Portions of the plait-point locus cannot be "seen" directly by the rich hydrocarbon gas because of curvature of the dew-point surface. In such a system, continuous injection of the rich gas over stationary reservoir fluid might form a zone of contiguously miscible compositions from pure rich gas to pure reservoir fluid by: (1) saturating the reservoir fluid with injected gas to the bubble-point surface; (2) creating by mass transfer with fresh injected gas a path of contiguously miscible compositions along the bubble - point surface to the plait-point locus; and (3) creating by mass transfer with additional injected gas a path of gas compositions along the dew-point surface up to the point where direct miscibility results between dew-point fluid and the injected rich gas. INTRODUCTION The use of the pseudo-ternary-phase diagram to illustrate miscible displacement phase behavior has been discussed by several authors.5,7 Such a representation of phase behavior is not rigorous, but the ternary diagram nevertheless gives a qualitative picture of what actually occurs in a miscible displacement process. Fig. 1 is a typical illustration of miscible displacement phase behavior by a ternary diagram. The multicomponent hydrocarbon system is divided into three pseudo-components: a light fraction containing methane and nitrogen, an intermediate fraction containing ethane through hexanes plus carbon dioxide, and a heavy fraction containing heptane and heavier components. A two-phase region is bounded by a dew-point curve and a bubble-point curve, which are joined at the critical point. The concept deduced from such a representation for miscible displacement by a condensing-gas-drive process is as follows: a rich gas G, which lies to the right of the limiting tie line through the critical point C, is injected into the reservoir and contacts reservoir fluid L, saturating the reservoir fluid to give bubble-point fluid L1 and equilibrium dew-point gas GI. Continued injection of rich gas changes the composition of the saturated liquid L1 through a series of liquid compositions lying along the bubble-point curve, until the critical composition C is reached, at which point direct miscibility with the rich gas is achieved. Some equilibrium gas with compositions lying along the dew-point curve from G1 to C is also formed in this process. Thus, continuous injection of rich gas forms a

Jan 1, 1966

By Leonard C. Grimshaw

The idea of bonding two dissimilar ferrous metals, and making use of both, is an old one. Tips have been brazed onto tool shanks for many years. The bonding of larger pieces to form whole bars and sheets has also been accomplished, but has received little attention in the technical literature, although a number of patents have been issued covering various methods for its accomplishment. There are three main motives behind the manufacture of composite steels. One is that the physical properties of one steel may supplement those of another; a useful, but at the same time, hard, brittle steel may be backed by a soft ductile one, or a stainless steel may be backed by a steel possessing greater strength. Another motive is direct lowering of cost; an expensive steel may be backed by a less expensive one, bringing down the cost of the piece as a whole. The third motive is an indirect lowering of cost, as when a useful steel is backed by a steel that is easier to machine, so that machining cost on a finished product is lowered considerably. Often more than one of these motives lead to a decision to use a composite steel. Previous Methods There are two methods in use that result in tolerably good bonds between ferrous metals: (1) the casting method, and (2) what we shall call the "heat-pressure" method. This "heat-pressure" method covers processes in which the bond is made between two prepared surfaces by heating them and then applying pressure. Usually sufficient pressure must be used to produce appreciable spread and elongation, or else we arc merely brazing and not truly welding. Casting methods for welding together ferrous metals have been used more than 74 years, and are described in the technical literature as applied

Jan 1, 1936