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By M. Berjak, D. M. Hawlins, R. B. MacGillivray

SYNOPSIS The paper describes a method of producing iso line plans using a mathematical technique for interpolating between boreholes, and its computer application. The iso line plans are used for assessing the value parameters of an ore body. The computer programme also produces means and standard deviations to supplement these iso line plans. The mathematical techniques tested for the production of a satisfactory fit of the data were the polynomial, Fourier and moving average technique. The iso line plans are used for assessing the quality of the ore and other salient features such as depth to the floor of the ore body at different points, with a view to deciding how best the ore may be mined. The plans are plotted by the computer to any scale required. INTRODUCTION To assess the economic exploitability of an ore body many tedious and repetitive calculations must be carried out. Observations of borehole or development samples have to be processed yielding measures of ore reserve and quality in each sampled area, and also measures of the worth of the ore body as a whole and of variation in quality from section to section of the ore body. The observations made on the ore (which we shall term 'value parameters') may embrace a variety of characteristics of the ore. Obvious value parameters are those directly related to the economics of mining, viz. assay values, ore body thickness and depth below surface. Other value parameters however may be almost as important. In the case of coal for example a purchasing contract may specify limits of acceptability on other value parameters such as calorific value, ash content or the percentage of volatile matter. In this case, a detailed study of these value parameters will also be crucial. This detailed study must include an assessment, not only of the body as a whole, but also the regional variation in value of the parameter. An obvious and satisfactory way of displaying this variation is by drawing iso lines of constant value parameter on a map of the ore body and supplementing this with the means and standard deviations of the value parameters taken over all boreholes. Such plans, considered together, show up the zones of significance. The relatively stereotyped nature of the operations involved in this process makes it amenable to computer solution. In order to produce iso line plans on the computer we require some mathematical formula for producing a 'valuation surface'-a mathematical

By M. D. Attfield

During 1992 a number of papers were published on aspects of respiratory disease in US underground coal miners. These included reports on trends in prevalence of coal workers' pneumoconiosis, the risks of developing coal workers' pneumoconiosis in relation to dust exposure, and the relationship between lung disease other than pneumoconiosis and dust exposure. The intent of this paper is to provide a summary of the essential findings from these investigations. PREVALENCE AND INCIDENCE OF COAL WORKERS' PNEUMOCONIOSIS Coal workers' pneumoconiosis (CWP) is usually detected in living miners through use of the chest x-ray. Deposits of dust in the lung, and the fibrotic tissue reactions to the dust, are manifested on the x-ray as small and large opacities. These abnormalities are quantified into degrees of severity using a standardized classification scheme (the International Classification of Radiographs of the Pneumoconioses of the International Labour Office). For small opacities, the scheme provides for a four-point scale of abnormality (0, 1, 2, 3), reflecting increasing profusion of the opacities on the x-ray. In this report, category 1 or greater small rounded opacities (CWP 1+) represents the earlier radiographic stages of pneumoconiosis. Large opacities, indicate the later and often disabling stage of pneumoconiosis, usually referred to as progressive massive fibrosis (PMF). Information collected in a national surveillance program for underground coal miners (the Coalworkers' X-ray Surveillance Program (CWXSP)) and a large epidemiologic study (National Study of Coalworkers' Pneumoconiosis (NSCWP)) provide a means to assess trends in CWP prevalence over the years following the reduction in dust levels mandated by the 1969 Federal Coal Mine Health and Safety Act (FCMHSA). These programs are operated by the National Institute for Occupational Safety and Health (NIOSH), and are described in detail in Attfield and Althouse (1992) and Attfield and Castellan (1992). Both studies were divided administratively into distinct time periods, and the results shown here reflect that division. Figure 1 shows the prevalence of CWP 1+ as determined from data from the CWXSP, plotted by time period and years in mining. There is evidence of a temporal reduction in prevalence. (The lack of an obvious decline in the early 1970s may be an artifact arising from use of different versions of the standard classification.) Figure 2 provides information in the same format for PMF, and shows a similar trend [ ]

Jan 1, 1993

Solid EnergyÆs project evaluation process utilises a standard staged gate approach, with five levels leading ultimately to project execution (level 6). For a project to progress through each level, specific financial criteria need to be satisfied. Solid Energy primarily use value based (discounted cash flow) metrics to support investment decisions. Within this approach Solid Energy utilise a number risk-based techniques to quantify project uncertainty. This includes the use of risk-based probabilistic analysis using Monte Carlo simulation and probability-based scenario analysis. Figure 1 shows the net present value (NPV) cumulative probability. Project risks are identified through comprehensive risk workshops. The output of these workshops (list of key risks, ranges of uncertainty and probability of occurrence), is quantified within the financial models. Incorporating this risk information on a probabilistic basis allows an investment decision to be made with a robust knowledge of the range of possible project outcomes along with the likelihood (probability) of those outcomes occurring, for example 80 per cent confidence that the project value will be greater than NZ$10 M, 50 per cent confident it will be greater than NZ$50 M and ten per cent confident it will be greater than NZ$70 M. Incorporating risk implicitly into the financial analysis allows the business to easily evaluate/measure projects with different risk profiles. The detailed risk analysis does away with the complexity of arbitrary risk-based hurdle (discount) rates, as projects can be evaluated based on their specific risk characteristics. Whether a project will progress to the next level of investment and ultimately into implementation depends firstly on the value assessment in comparison with the associated risk/uncertainty, and secondly matching this with the businesses appetite for risk. At Solid Energy standard criteria are used for specific project evaluations, but in addition projects are considered on a portfolio or strategic basis also. Solid Energy attempts to maintain a mix of low risk (low value) projects with some high risk (high value) opportunities. The mix of projects must remain in line with the shareholdersÆ risk appetite. One of the keys to a successful investment decision is ensuring that the appropriate information is available to decision-makers so that they may proceed with a comprehensive understanding of the risks and impacts associated with the decision.

Jan 1, 2007

By L. O. Howard

WHAT was perhaps the first attempt at basic converting in the Southwest was made by the late Charles F. Shelby at Cananea early in 1907, when he removed the acid lining from one of the 8 by 12-ft. barrel-type converters then in use at the reduction works of the Cananea Consolidated Copper Co. and substituted a lining of one course of magnesite brick. In this shell, he blew 21 taps (about 180 tons) of matte to white metal before the brick gave out along the tuyère line. In every case, the white metal so made was transferred and blown to copper in other converters. The cost of the lining was $700 and Mr. Shelby gave up the experiment as impracticable. No further serious effort to bessemerize copper mattes in a vessel having a more or less permanent lining of basic brick was made by any of the large copper smelters in the Southwest until some months after Messrs. Peirce and Smith had brought the question permanently before the public, and had proved it to be a metallurgical and financial success by their work at Baltimore and Garfield. Early in 1911, the Cananea Consolidated Copper Co. had in operation at its reduction works at Cananea, Mex., several 8 by 12-ft. barrel-type converters lined with magnesite brick which had previously been operating with acid linings. In the 18 months following, these converters produced 50,000 tons of copper. In June, 1912, this company installed the first of its new equipment of six stands of electrically operated 12-ft. Great Falls type converters, and shortly afterward discontinued the use of the smaller shells. In September, 1911, the Consolidated Kansas City Smelting & Refining Co. had installed at its El Paso plant two 10 by 26-ft. Peirce-Smith converters. Later, in March, 1913, one standard Great Falls type converter was added. .About the time the Peirce-Smith converters were going in at El Paso, the Calumet & Arizona Mining Co. was experimenting with basic linings in 7-ft. by 10-ft. 6-in. barrel converters at its plant in Douglas, Ariz. These efforts were sufficiently successful to enable the plant to handle the total output of copper in these small shells until the standard. Great Falls converters were substituted in June, 1914.

Jan 9, 1916

By R. D. Hamilton, W. J. Miles

In 1987, the International Agency for Research on Cancer designated crystalline silica as a 2A probable carcinogen. As a result, crystalline silica is regulated under the Hazard Communication Standard as a probable carcinogen at a concentration of 0.1% or greater in the workplace. Crystalline silica will be listed as "reasonably anticipated to be a carcinogen" by the National Toxicological Program. This designation of crystalline silica as a probable carcinogen has created major issues for the mineral and chemical producer, including labeling, hazard communications, marketing a material with potential hazard exposure, and potential liability. The State of California also regulates crystalline silica under Proposition 65 in products as a carcinogen, based upon the IARC determination that it is a probable carcinogen. More federal and state regulations are anticipated as a result of the IARC classification of crystalline silica and its subsequent federal regulation. Respirable crystalline silica has been regulated as a hazardous dust for many years by OSHA. This regulation was initiated to prevent silicosis, another occupational disease. OSHA has established monitoring devices for determining the concentration of airborne respirable silica in order to measure exposure to this substance. The accurate determination of the crystalline silica content is critical to dealing with these health risks and regulations. In the past, crystalline silica was considered to be safe and, because of its ubiquitous nature, little effort was made to determine or quantify its presence in most materials. The Hazard Communication Standard requires that a1l products containing 0.1% or greater crystalline silica must be properly labeled for this material and its hazard. Unfortunately, the 0.1% threshold was not chosen on the basis of established analytical capabilities for crystalline silica, but results from general limits for potentially carcinogenic chemicals and substances. Numerous methods have been developed for the determination of crystalline silica in a variety of products. However, these methods are not generally applicable since they must be specific for silica in any of its crystalline states and not just for total silica content. Chemical tests for silica are not difficult and quite accurate at the 0.1% level; however, in addition to its crystalline phases, it occurs in amorphous phases and in a wide variety of crystalline and amorphous silicates. In general, there is little correlation between total silica content and crystalline silica.

Jan 1, 1991

By L. O. Howard

What was perhaps the first attempt at basic converting in the Southwest was made by the late Charles F. Shelby at Cananea early in 1907, when he removed the acid lining from one of the 8 by 12-ft. barrel-type converters then in use at the reduction works of the Cananea Consolidated Copper Co. and substituted a lining of one course of magnesite brick. In this shell, he blew 21 taps (about 180 tons) of matte to white metal before the brick gave out along the tuyhre line. In every case, the white metal so made was transferred and blown to copper in other converters. The cost of the lining was $700 and Mr. Shelby gave up the experiment as impracticable. No further serious effort to bessemerize copper mattes in a vessel having a more or less permanent lining of basic brick was made by any of the large copper smelters in the Southwest until some months after Messrs. Peirce and Smith had brought the question permanently before the public, and had proved it to be a metallurgical and financial success by their work at Baltimore and Garfield. Early in 1911, the Cananea Consolidated Copper Co. had in operation at its reduction works at Cananea, Mex., several 8 by 12-ft. barrel-type converters lined with magnesite brick which had previously been operating with acid linings. In the 18 months following, these converters produced 50,000 tons of copper. In June, 1912, this company installed the first of its new equipment of six stands of electrically operated 12-ft. Great Palls type converters, and shortly afterward discontinued the use of the smaller shells. In September, 1911, the Consolidated Kansas City Smelting & Refining Co. had installed at its El Paso plant two 10 by 26-ft. Peirce-Smith converters. Later, in March, 1913, one standard Great Falls type converter was added. About the time the Peirce-Smith converters were going in at El Paso, the Calumet & Arizona Mining Co. was experimenting with basic linings in 7-ft. by 10-ft. 6-in. barrel converters at its plant in Douglas, Ariz. These efforts were sufficiently successful to enable the plant to "handle the total output of copper in these small shells until the standard Great Falls converters were substituted in June, .1914.

Jan 1, 1917

By B. Gamble

Solid Energy's project evaluation process utilises a standard staged gate approach, with five levels leading ultimately to project execution (level 6). For a project to progress through each level, specific financial criteria need to be satisfied. Solid Energy primarily use value based (discounted cash flow) metrics to support investment decisions. Within this approach Solid Energy utilise a number risk-based techniques to quantify project uncertainty. This includes the use of risk-based probabilistic analysis using Monte Carlo simulation and probability-based scenario analysis. Figure 1 shows the net present value (NPV) cumulative probability. Project risks are identified through comprehensive risk workshops. The output of these workshops (list of key risks, ranges of uncertainty and probability of occurrence), is quantified within the financial models. Incorporating this risk information on a probabilistic basis allows an investment decision to be made with a robust knowledge of the range of possible project outcomes along with the likelihood (probability) of those outcomes occurring, for example 80 per cent confidence that the project value will be greater than NZ$10 M, 50 per cent confident it will be greater than NZ$50 M and ten per cent confident it will be greater than NZ$70 M. Incorporating risk implicitly into the financial analysis allows the business to easily evaluate/measure projects with different risk profiles. The detailed risk analysis does away with the complexity of arbitrary risk-based hurdle (discount) rates, as projects can be evaluated based on their specific risk characteristics. Whether a project will progress to the next level of investment and ultimately into implementation depends firstly on the value assessment in comparison with the associated risk/uncertainty, and secondly matching this with the businesses appetite for risk. At Solid Energy standard criteria are used for specific project evaluations, but in addition projects are considered on a portfolio or strategic basis also. Solid Energy attempts to maintain a mix of low risk (low value) projects with some high risk (high value) opportunities. The mix of projects must remain in line with the shareholders' risk appetite. One of the keys to a successful investment decision is ensuring that the appropriate information is available to decision-makers so that they may proceed with a comprehensive understanding of the risks and impacts associated with the decision.

Jan 1, 2007

By Frank L. Antisell

Certain physical and chemical properties of copper are so intimately related that a change in variation of the physical properties indicates a certain chemical change. The standard specifications of copper of the American Society for Testing Materials are based upon physical characteristics with one exception, namely, the copper contents. The physical standard includes electrical properties, defects in the "set or casting," variation in dimensions, weight, etc. Paragraph 4 of the specifications states, " Wire bars, cakes, slabs and billets shall be substantially free from shrink holes, cold sets, pits, sloppy edges, concave tops, and similar defects in set or casting." Shrink holes and concave tops indicate both physical and metallurgical defects; cold sets, sloppy edges, etc., indicate physical defects only; small pit holes may be indicative of either or both physical and metallurgical defects, depending on conditions under which they originated. To the refiner, a copper casting, no matter how made, is known as a "shape," irrespective of whether it is wire bar, ingot, cake, slab, billet, etc. As practically all shapes are cast in an open mold, that is a mold without cope, the surface of the shape cools in contact with the atmosphere and has a coating of oxide. Certain phenomena of the surface indicate the adaptability of the metal for forging or rolling. The writer believes that it is much simpler to produce shapes that are metallurgically correct than those that are mechanically satisfactory. When copper is to be rolled into thin sheets or drawn into very fine wire, the slightest physical defect becomes evident. It is, therefore, of the utmost importance to consider every characteristic of the shape and ascertain the physical and chemical relationship. Pitch and Set.—The appearance of the surface of the copper when cooled is considered as to its "pitch," or the general contour of the surface of the shape (which may vary from a concave to a convex); and the '" set,"' or the detail appearance of the surface of the shape (the wave-like structure). When the shape presents a concave surface, it is known as "low pitch;" a convex surface is known as "high pitch" or "tough pitch;" and a level surface is called "flat pitch." The set of the copper is, in a very large measure, directly related to the pitch; and so intimately are

Jan 1, 1921

By Robert E. Green, Edmund G. Henneke

The velocities of the three possible modes of elastic wave propagation have been calculated for single-crystal copper at 1-deg intervals throughout the standard stereographic triangle. The results are presented as isospeed contours. The deviations between the direction of maximum energy flux and specimen axes are presented as arrows drawn on stereographic projections. The orientation factors associated with dislocation motion on the primary slip system only and also on all twelve slip systems have been calculated for all three elastic waves. The orientation factors associated with a longitudinal standing wave resonant specimen have also been calculated. The results of these calculatians are compared with earlier calculations on aluminum crystals and the relevance of these results to ultrasonic attenuation measurements is discussed. GREEN and inton' have recently outlined a procedure for determining the orientation factors for dislocation damping in single crystals. The stress field associated with a given mode of elastic wave propagation is resolved on the slip systems in which it is assumed the dislocations are vibrating due to this stress field. Since, at room temperature, slip has been observed on a limited number of systems in fcc and hcp metal crystals, it is assumed that these systems are the appropriate ones to use for the resolving process. This procedure has been applied to aluminum crystals by Hinton and hreen' and to zinc crystals by Henneke and Green.3 In the present paper the modes of elastic wave propagation have been calculated for copper crystals possessing orientations at 1-deg intervals throughout the standard stereographic triangle. The deviation of the energy flux vectors from the specimen axis have been calculated for each mode. The stress field associated with each mode has been resolved on the (111)(110) slip systems in order to determine the orientation factors for dislocation damping. MODES OF ELASTIC WAVE PROPAGATION For any direction in a linear homogeneous aniso-tropic material three elastic modes can be propagated. In general, each of these modes has both longitudinal and transverse particle displacements associated with it and thus are not "pure" modes. However, there is always one component which is greatest in either a longitudinal or transverse direction and hence the mode may be referred to as quasi-longitudinal or quasi-transverse, accordingly. For certain axes of high symmetry pure modes may be propagated. The phenomenon of refraction of the energy flux vector is also exhibited by elastic wave propagation in ani-

Jan 1, 1969

By A. J. Lanza

The prevention of illness among the employees of the mining industry is especially important in view of the importance of the industry, the unsettled conditions of labor, which emphasize the economic necessity for conservation of labor, and for humanitarian reasons. The burden of chronic illness is altogether unwarranted and unnecessary. If the industry is to maintain its labor supply in an efficient manner and keep the standard high, it is essential that it make every effort to keep its labor in good physical condition. The basic principles underlying health conservation are the same for any industry, but they are here applied especially to the mining industry, and more particularly, perhaps, to metal mining, as it is with this phase of the industry that the writer is most familiar, and as it is conceded that metal mining presents more health hazards than coal mining. The first step in the prevention of illness lies in securing employees who are in sound health and free from organic disease. I do not mean that all men who go underground should be physically perfect, but they should be free from organic defects of the heart, lungs, and other organs, or from anatomical defects that would markedly increase their liability to accident. The number of men working underground in mines in this country whose physical condition totally unfits them for such work presents a situation that should not be allowed to continue, and which I believe is not equalled in any other industry in the country. There is but one way to secure men who come up to the required standard, and that is by a thorough physical examination before employment. I am fully aware of the objections felt toward such a procedure by many employers and employees. There is much to be said on both sides, but to any one who is at all familiar with the mining industry as it is today in this country, and who approaches the matter with an unbiased mind, there can be absolutely no doubt as to the necessity for physical examination. As a purely logical proposition this is self-evident, but there remains the further necessity of placing physical examinations on such a basis that they can be administered in a just and equitable manner free from the abuses and suspicions to which they have been subject in the past.

Jan 1, 1919

In 1953, Dr N Fisher of the Bureau of Mineral Resources, acting on behalf of the Society of Economic Geologists, which is considering the sponsoring of a uniform system of nomenclature for ore reserves, enquired from The Australasian Institute of Mining and Metallurgy whether it had ever considered this subject and particularly whether it had ever set up any standard definitions or specifications with regard to ore reserves.As this had not been done, but was thought to be desirable, the Council of The Institute appointed a Committee consisting of H H Dunkin, Associate Professor of Mining Engineering, University of Melbourne (Convenor); E A Rudd, Professor of Economic Geology, University of Adelaide; J D Campbell, Chief Geologist of Western Mining Corporation; and the writer, to consider the report on this subject.In December 1953, the Committee submitted its frrst report to the Council and this report (reproduced here as an Appendix) was widely circulated among member companies and engineers, with an invitation to comment on the proposals. A large number of replies was received and studied.After studying the replies the Committee issued a second report, in part as follows:From these replies we have learned that the suggestions made in the frrst report of the Committee are unacceptable to most of those interested, on one or all of the following grounds:a. number of categoriesb. defmition of categoriesc. nomenclature.It is also apparent that all alternatives suggested in the various replies would likewise be unacceptable to the majority of those replying.As a result, we have come to the conclusion that neither this Committee nor any other could at present and in the present state of opinion on the subject make recommendations regarding classification and nomenclature of ore reserves which would be acceptable to a sufficiently large proportion of members.It seems to us now that the objective of a widely applicable and acceptable system of classification and nomenclature, one that could be endorsed by The Institute and ultimately become standard practice in the industry, requires more than compromise between present conflicting views... it requires first a foundation of substantial agreement on the essential facts, implications and range of ore reserve estimation. This does not yet exist and it is evident that much additional analysis of present practice and opinion must be undertaken before any further systhesis is attempted.

Jan 1, 1994

By James E. Harding

The usual method of sampling mineral deposits is to drill holes and assay the sludge or core. Though the results thus obtained may not represent the true average value of the deposit, it is on these results that estimates of practically all large orebodies are made. In a large number of cases, the results of exploitation fall far short of the estimated value of the orebody. The standard method of making these estimates is to find the cubic contents of triangular prisms, in the apexes of which the drill holes are placed, by. multiplying the surface area by the average depth of the ore in the three drill holes and then multiplying the cubic contents by the specific gravity of the ore or rock to find the tonnage. The assay value is found by dividing the sum of the products of the depth of ore in each hole and the corresponding assay value by the sum of the depths of the ore in the three holes. As the latter part of this method is subject to many mathematical errors, it should not be depended on. Some engineers, therefore, use a discount factor, but as this factor is either arbitrarily selected or is obtained by obscure methods, the results are no better than guesswork. The standard method is correct when the drill holes are so laid out that the triangular prisms are equilateral in cross-section. But as ore deposits cannot always be drilled into at regular intervals and because of the human factor, if many holes are drilled, triangular prisms of all degrees of angularity are produced. In such cases it is necessary to assume that the influence of the different holes is not the same in all directions, which assumption of course is absurd. The area of the triangular prism is usually found by scaling; an ore-body of sufficient size to justify churn drilling is too large to be handled conveniently on maps having a scale larger than from 1000 to 1 to 1,500 to 1 and maps of such scales cannot be accurately measured. A variable is thus introduced in the first operation. In addition, the scaling is frequently done from blueprints, which sometimes shrink as much as 4 per cent. between printing, washing, and drying. The error due to scaling may be eliminated by surveying the location of each hole and

Jan 1, 1922

By Alexander Squire

ONE of the principal factors that have contributed to the hesitancy of design engineers to use metal-powder parts is the difficulty experienced in the determination of the mechanical properties of compacted components. Because standard test: methods cannot usually be applied in the inspection of such parts, and because adequate special test methods have not been developed, it has been considered desirable to investigate the methods that are applicable. These may then be included in specifications for metal-powder materials' and components in order that the design engineer may be assured that the subject parts possess the desired mechanical properties for satisfactory performance. It is known that direct relationships exist between usually nonmeasurable properties as tensile strength and elongation and shear strength and the angle through which compacts may be plastically deformed before fracturing. These correlations are valuable in the inspection of specific parts, but it is necessary to set up standards before they may be satisfactorily employed. Apart from the use of correlations, the number and character of the inspection methods that are applicable to almost all powder parts is limited to three-chemical analysis, density, and hardness. The last of these may or may' not be a pertinent characteristic, since the indentation hardness of many powdered-metal materials is so low as to have little significance. It is, however, occasionally desirable to use hardness measurements as an indication of the uniformity of the structure of pieces of a given type. Chemical analysis reveals very little about the physical properties of metal-powder compacts, since the unpressed powder is usually of a composition almost identical with that of the finished pieces. Density determinations are simple, to perform and have the advantage of being nondestructive in character. If the dimensions of a machine part are known (and usually they are), the average density may be determined by weighing a given number of pieces and dividing by the calculated volume of the parts. More accurate measurements may be made by the water-displacement method. , It. is recognized that mechanical properties of metal-powder materials are related to their density, but usually these relations are determined for only one type of powder, processed under a given set of conditions. Comparisons of the properties of compacts formed from different types of powders are made on the basis of compacting... pressures instead of density, and for. this reason the results obtained by using powders of different types appear widely variant.

Jan 1, 1947

By Alexander Squire

One of the principal factors that have contributed to the hesitancy of design engineers to use metal-powder parts is the difficulty experienced in the determination of the mechanical properties of compacted components. Because standard test methods cannot usually be applied in the inspection of such parts, and because adequate special test methods have not been developed, it has been considered desirable to investigate the methods that are applicable. These may then be included in specifications for metal-powder materials and components in order that the design engineer may be assured that the subject parts possess the desired mechanical properties for satisfactory performance. It is known that direct relationships exist between usually nonmeasurable properties as tensile strength and elongation and shear strength and the angle through which compacts may be plastically deformed before fracturing. These correlations are valuable in the inspection of specific parts, but it is necessary to set up standards before they may be satisfactorily employed. Apart from the use of correlations, the number and character of the inspection methods that are applicable to almost all powder parts is limited to three—chemical analysis, density, and hardness. The last of these may or may not be a pertinent characteristic, since the indentation hardness of many powdeied-metal materials is so low as to have little significance. It is, however, occasionally desirable to use hardness measurements as an indication of the uniformity of the structure of pieces of a given type. Chemical analysis reveals very little about the physical properties of metal-powder compacts, since the unpressed powder is usually of a composition almost identical with that of the finished pieces. Density determinations are simple to perform and have the advantage of being nondestructive in character. If the dimensions of a machine part are known (and usually they are), the average density may be determined by weighing a given number of pieces and dividing by the calculated volume of the parts. More accurate measurements may be made by the water-displacement method. It is recognized that mechanical properties of metal-powder materials are related to their density, but usually these relations are determined for only one type of powder, processed under a given set of conditions. Comparisons of the pioperties of compacts formed from different types of powders are made on the basis of compacting pressures instead of density, and for this reason the results obtained by using powders of different types appear widely variant.

Jan 1, 1947

By Alexander Squire

One of the principal factors that have contributed to the hesitancy of design engineers to use metal-powder parts is the difficulty experienced in the determination of the mechanical properties of compacted components. Because standard test methods cannot usually be applied in the inspection of such parts, and because adequate special test methods have not been developed, it has been considered desirable to investigate the methods that are applicable. These may then be included in specifications for metal-powder materials and components in order that the design engineer may be assured that the subject parts possess the desired mechanical properties for satisfactory performance. It is known that direct relationships exist between usually nonmeasurable properties as tensile strength and elongation and shear strength and the angle through which compacts may be plastically deformed before fracturing. These correlations are valuable in the inspection of specific parts, but it is necessary to set up standards before they may be satisfactorily employed. Apart from the use of correlations, the number and character of the inspection methods that are applicable to almost all powder parts is limited to three—chemical analysis, density, and hardness. The last of these may or may not be a pertinent characteristic, since the indentation hardness of many powdeied-metal materials is so low as to have little significance. It is, however, occasionally desirable to use hardness measurements as an indication of the uniformity of the structure of pieces of a given type. Chemical analysis reveals very little about the physical properties of metal-powder compacts, since the unpressed powder is usually of a composition almost identical with that of the finished pieces. Density determinations are simple to perform and have the advantage of being nondestructive in character. If the dimensions of a machine part are known (and usually they are), the average density may be determined by weighing a given number of pieces and dividing by the calculated volume of the parts. More accurate measurements may be made by the water-displacement method. It is recognized that mechanical properties of metal-powder materials are related to their density, but usually these relations are determined for only one type of powder, processed under a given set of conditions. Comparisons of the pioperties of compacts formed from different types of powders are made on the basis of compacting pressures instead of density, and for this reason the results obtained by using powders of different types appear widely variant.

Jan 1, 1947

By Hallan N. Marsh

The subject of mud sounds so simple, uninteresting and unimportant that it has failed to receive the attention that it deserves, at least as applied to the drilling of oil wells. As a matter of fact, it is one of the most complicated, technical, important and interesting subjects in connection with rotary drilling. In 1923, R. E. Colloml said "It should be possible to establish certain physical standards for mud fluid," but while considerable work has probably been done on the subject before and since that time, and numerous articles have been written, the progress of general knowledge of the subject and the application of better practices have not kept pace with other developments in rotary drilling, and physical standards cannot be considered as established. Mud is apt to be thought of as an accessible and cheap material, but its cost is likely to be an important factor in the total cost of drilling a well. Some wells "make" most of their own mud out of the formations drilled, but in most cases a great deal of mud-making material has to be supplied from more or less distant sources. Expenditures for mud materials by one California company totaled practically one-quarter million dollars during 1929 and averaged about $13 per rig per drilling day, or about $2000 per well. Under unfavorable conditions, mud cost may reach many times this figure. Parker2 says that in the Hobbs, N. M., field "as much as $75,000 has been spent on one well alone for mud and mud-weighting compounds."' Presumably this is an exceptional case. Unit cost of mud fluid will range from 10 c. per barrel, where suitable material is close at hand, to $1 where crude material is available at moderate price but must be hauled long distances, and on up to as much as $11 per barrel where high gas pressure requires the use of special processed weighting material. The cost of mud pales into insignificance, however, when the results of using improper mud are considered. Blow-outs and lost circulation are obviously the result of inadequate mud. Stuck drill pipe, whether or not twisted off, in many if not most cases probably is caused by cuttings settling out of the mud or the mud itself precipitating. Recent experi-

Jan 1, 1931

By L. G. Austin, P. T. Luckie, R. R. Klimpel, R. S. C. Rogers

1. INTRODUCTION The Bond method(1.2) for sizing tumbling ball mills has been used successfully for many years. The primary purpose of the Bond calculation is to predict the mill size and mill power to give a desired capacity (Q tph) from a wet overflow ball mill in normal closed circuit operated with a circulation ratio of 2.5. The input to the calculation is the make-up feed size characterized by the 80%-passing size in micrometers (xG), the desired circuit product size characterized by the 80%-passing size (xQ), and an empirical grindability number (the standard Bond Work Index) determined by a standard test. A series of empirical correcting factors(2) are used to convert the results to open-circuit or dry grinding in grate discharge mills. The Bond method has two major engineering advantages: (i) it is very simple; (ii) experience has shown that it works for many circumstances, as a reasonable approximation, because it is based on direct conversion of test data to results measured on industrial mills. There are four main disadvantages of the Bond method. First, there will be a range of error for any specific mill and material and set of operating conditions because the method does not explicitly include several variables which are obviously important: (i) recycle ratio and classifier efficiency (ii) mixture of ball sizes in the mill (iii) variations of residence time distributions with mill geometry and slurry density (iv) lifter design (v) slurry rheology (vi) different underfilling or overfilling of the mill as flow rate is changed. Second, the Bond method explicitly assumes that specific grinding energy E is not a function of ball load, whereas industrial practice and laboratory tests(3) show that the specific grinding energy (to go from a specified feed to a specified product) is less for lower ball loadings than the ball loading for maximum mill capacity. Third, the method uses only the 80% passing size of circuit feed and product to characterize the size distributions, whereas mill capacity in general must depend on the shape of the feed size distribution and the product size distribution. The prime example of this is the use of the reverse closed circuit which removes some fraction of the make-up feed before it is

Jan 1, 1982

By G. B. Fritts, A. C. Moschetti

"INTRODUCTION The act of Congress establishing the Bureau of Mines in 1910 delegated to the Bureau, as one of its principal duties and functions, the study of health and safety in the mineral industries with a view to devising and putting into operation means of conserving the health and safety of persons employed in the industries. One of the first health and safety activities of the Bureau of Mines was therefore organization of a trained force to promote instruction in first aid in the mineral industries. First-aid training has been an important activity of the Bureau of Mines since its establishment. The standard first-aid course requires a minimum of 15 hours of instruction and practice work. A person completing the initial training course is issued an engraved certificate and pocket card certificate; thereafter, only a pocket card certificate is issued for additional training.From 1910 to 1926 first aid was taught by trained Bureau of Mines employees; consequently, only a comparatively few persons were trained. In 1926, however, cooperative first-aid training was initiated, which provided the means to train the entire personnel of mines or plants within a short time in a single campaign. Many mines and plants have taken advantage of the cooperative plan, whereby keymen in a mine or plant, in addition to the initial training received, are given intensive training in first aid by a Bureau of Mines instructor. Those men who show an aptitude for teaching are selected as tentative provisional first-aid instructors. In order to qualify as a provisional instructor, a tentative instructor must, under the supervision of a Bureau of Mines instructor, train a class of at least six persons employed in the minerals industries in the standard Bureau of Mines course in first aid. After it has been determined that the class has been trained properly, a provisional first-aid instructor's pocket card certificate is issued to the class instructor. The first-aid instructor can renew his pocket card certificate each year by training or retraining at least six persons connected with the mineral industries to satisfactorily pass an examination given by a Bureau of Mines representative. After five yearly renewals of a provisional instructor's pocket card certificate, an engraved, permanent instructor's certificate and a permanent instructor's pocket card certificate are issued to the instructor."

Feb 1, 1955

By C. A. Herbert, C. M. Young

WithIn the past few years, considerable development has been made in the coal-mining industry in Illinois and Indiana and it is the purpose of the authors to record its most important phases. Perhaps the two most striking features are the entry into the producing fields of certain large consumers of coal and the magnitude of some of the new operations. Mines are now being equipped by the Chicago & Northwestern R. R., the Chicago, Burlington & Quincy R. R., the Standard Oil Co., of Indiana, and the Union Electric Light & Power Co. of St. Louis. Besides, the U. S. Steel Corpn. has increased its coal-mining activities by commencing operations on its large holdings northeast of Benton, Franklin CO., Ill. Until recent years, the supply of coal has commonly been adequate and prices have generally been favorable, especially to large consumers, as the magnitude of their purchases enabled them to obtain satisfactory quotations from producers. In fact, most consumers of coal have believed that they could buy their fuel more satisfactorily in the open market than they could produce it themselves. The entry of the United States into the war was accompanied by various disarrangements of industrial conditions, and a demand for coal in excess of the supply seemed likely to be experienced for an indefinite period. Under these circumstances it seemed that the greatest assurance of a supply of fuel was its production by the operation of mines. In the case of the Chicago & Northwestern R. R. and the U. S. Steel Corpn. the production of coal has been carried on for many years with satisfactory results, but the C. B. & Q. R. R., the Union Electric Light & Power Co. and the Standard Oil Co. of Indiana had not been producers of coal, at least in this district. Apparently the most impressive feature of the newest mines is their capacity, for in some cases it is planned that production shall be in the neighborhood of 1000 tons per hr. There are mines now in this field having nearly that output, but these were not planned for such output; in fact, when the shafts were sunk and the machinery installed it was generally supposed that the daily outputs would be only in the neighborhood of 4000 to 5000 tons. Incidentally, there has risen

Jan 1, 1920

By J. M. Goris

"Researchers at the U.S. Bureau of Mines are conducting research on cable-type ground supports to assess their material and support properties and to provide design criteria for implementing cable supports as a workable roof control system under various underground mining conditions. Much of this work involves laboratory assessment of the material and support properties of cable supports using both conventional and birdcage steel cables. Included in this assessment are single and double cables, effects of changes in embedment length on support strengths, water-cement ratios, use of breather tubes, and pumpability and water-bleeding properties of grouts.IntroductionSince cable supports were introduced in the early 1970s, considerable research has been conducted in laboratories and at mine sites to understand their behaviour. The U.S. Bureau of Mines began a laboratory evaluation project in 1985 at the Spokane Research Center (SRC) to study the behaviour of cables supports under uniaxial loads so that the effects of various cable support components, such as grout and breather tubes, could be identified. Other objectives were to compare the behaviour of various types of cables and grouts and to collect meaningful data for numerical analyses. Pull tests were conducted to determine how cables supports reacted under uniaxial loads and how they failed when ultimate loads were exceeded. Such data are essential for the development of numerical models for analyzing rock mass behaviour when the rock is supported with cables.Test ProgramBecause each cable support can be composed of either single or double cables as well as a breather tube, pull tests were designed to include a variety of samples containing these components. A series of samples containing a single conventional cable, a neat cement grout having a water-cement ratio of 0.45, and no chemical additives was selected as a standard (Table I). Then one component in the system, such as the presence of a breather tube, was changed, and results of tests compared to results obtained for the standard samples to determine what influence that component might have on the behaviour of the support system. If significant changes in test data occurred, they could be attributed to that one component. To control the quality of the samples, the physical condition of the cables and tubes were monitored closely, and compressive and tensile strengths as well as flow properties were determined for the grout used in each test series."

Jan 1, 1991