Search Documents

By T. P. Meloy

In a recent paper, Gaudin and Meloy1 derived a size distribution function for single, impact fracture. Mular2 extended the use of this size distribution to describe the output of carefully controlled grinding experiments. As it stands, the function is difficult to compute. Consequently, a series of tables of the function are presented below: where M(x) is the cumulative fraction finer than size x when x is particle size. Weight percents are tabulated for the single parameter, r, in three different size ranges: from 0 to 1, from 1 to 10 and from 10 to 100. Tables I, II and III are cumulative percent finer and Tables IV, V and VI are for weight percent passing screen size x, retained on the screen below. A number of conditions from the table are plotted in Figs. 1 and 2. Fig. 1 is a plot of the cumulative percent finer vs. size, and Fig. 2 is the weight per-

Jan 1, 1963

By James R. Evans

Forecasting demand is an art as well as a science, and much personal judgment is required. National forecasts made for sand and gravel, crushed stone, and/or aggregate may be misleading or unhelpful locally. This is because of significant variation in demand within different geographic and/or political areas in the US, and/or the use of an inadequate data base and forecasting methodology. The best forecasts may be made for local or regional market areas, irrespective of geographic or political boundaries, provided a proper data base is available.1 The Bureau of Mines compiles detailed statistics on commodities sold or used in the construction industries and, therefore, is in an advantageous position to make national aggregate demand forecasts.

Jan 1, 1979

By R. D. Lord

The search for the best practical means of storing sulfide bearing tailings, where there is no residual excess of carbonate material is discussed in this paper• Usually the sulfide content decomposes, with the aid of bacterial action, and the resulting sulfuric acid escapes, along with any heavy-metal solutes, through embankments that are usually porous to some degree• The problem is typified in the tailings of the uranium operations of Elliot Lake, Ont., where mining started some 20 years ago• The approach to tailings disposal paralleled the practice for other hydrometallurgical plants treating gold and base-metal ores• Impoundment areas were designed to retain solids, and a clear and neutral overflow was considered satisfactory practice• Now experience has shown that these areas, some of which have been idle for over a dozen years, release acids in seepage and overflows to an unacceptable degree• To protect natural water courses, neutralizing plants are operated wherever required• Lime slurry is fed continuously into the tailings outflows in a quantity sufficient to raise the pH to 8•5 and precipitate heavy metals that may be in solution• The objection to this procedure is that the plants will require servicing indefinitely, unless a better remedy is found• The problem differs only slightly from that common to base-metal concentrators in that here the ore has been leached with sulfuric acid for the recovery of uranium• Any native content of calcareous material has been digested, and only that added for final neutralization is available to maintain a pH unfavorable to bacterial activity• Chemical oxidation slowly lowers the pH and when this reaches a level of 4•5 or less, bacteria become active and greatly accelerate the formation of acid. The bacterial process is probably at least ten times as fast as the chemical oxidation• Location and Processing The operations referred to, uranium and one copper mine, are located at approximately 46°N and 82°W longitude• This is typical Canadian Shield country, a land of lakes, deeply glaciated and rocky, with sparse soil which supports mixed forest cover• Drainage is to Lake Huron, 25 miles to the south• Average temperature is 45°F, ranging from -40° to +95°F• Annual precipitation is 38 in•, about half of which is snow• The ore is Precambrian, quartz-pebble conglomerate, with mineralization in the matrix• From 5 to 10% pyrite is present• All known means of pre-concentration have been tested, but a bulk sulfuric acid leach has proved the most efficient. Tailings have from the outset been neutralized before release• Current practice is to add ground limestone to bring the pH to 4•5, and then lime to raise the value to 10•5• Environmental regulations have recently been increased and the foregoing meets the new standards• Separate measures are taken to precipitate radium• Remedial Measures Since the outstanding environmental problem is the oxidation of pyrite by bacterial action, the solution is to contain the products, or arrest the process• Given the ambient temperature, favorable half of the time, four items are essential to the activity• 1) Pyrite• 2) Moisture pH < 4•5. 3) Oxygen• 4) Bacteria• Removing any one of these out of the range of tolerance will bring the reactions under control• A variety of proposals considered, and a number tested for the arrest of the process, are: (a) render embankments impermeable, (b) provide an impermeable cover, (c) cover with an oxygen absorbing layer, (d) provide a vegetative cover, (e) flood the site, (f) remove pyrite from current tailings, (g) add excess limestone to current tailings, (h) poison the bacteria• Bank Seal-On existing impoundment areas, where the embankments are several thousand yards in length, it is believed that any program of injecting sealants can have small chance of success• However, a moisture barrier is an indicated specification for future construction, and this can be highly expensive• Surface Seal-Depending on the configuration of the deposit, the downward travel of water should be prevented, and oxygen excluded• Burying a plastic membrane just below the surface has been considered, as has the application of a liquid sealant that would penetrate the surface. The objection to these remedies is the excessive cost of dealing with large areas and the expectation of only temporary benefit as a result• Frost penetration is over 4 ft, and frost action breaks up asphalt paving and all but heavy concrete in a few years• Organic Layer-An oxygen-absorbing layer, such as bark fines from paper mills has been proposed as a surface treatment• Cultivated into the tailings such material might be expected to arrest subsurface oxidation for some years• Estimates are 100 tons per acre of bark fines, or 35 tons per acre of sawdust, and these enormous quantities do not so far give assurance of providing a long-term remedy• Vegatative Cover-Several obvious benefits would result from a good growth of grass or other vegetation on abandoned tailings• While restoring the natural green of the tract the growth would prevent wind-blown dust and reduce erosion• Subsurface oxidation should be reduced, as well as the upward movement of ground moisture as occurs in dry weather. To this end, considerable research and field testing has been carried out to arrive at a formula - a prescription which will provide a self-sustaining growth on the tailings surface, or at least one that would survive with reasonable maintenance attention. Many test plots have been run with different combinations of surface treatment and seed mixtures. Generally, by addition and close cultivation of limestone, lime, and fertilizers, technical success has been demonstrated• Plants with a high tolerance for acid soil seem the more hardy, and a pH above 3 is indicated so that nutrients can be absorbed• Recommendations are for 12 to 15 tons of

Jan 1, 1977

By R. H. Heinen

Introduction Drilling is the initial operating step In open pit mining. It goes hand-in-hand with the blasting operations to ensure adequately broken material for the excavation equipment employed. The drilling effectiveness is highly dependent on the quality of the drill evaluation. This discussion will center on the parameters that are needed to evaluate drill requirements for a new mine The procedure that is generated can, however, be used to evaluate whether to expand existing operations and replace existing equipment. Methods of breaking ground, other than blasthole drilling and blasting, such as ripping, will not be discussed. The importance of correct drill evaluation extends to the other facets of open pit mining. Inadequate fragmentation, the direct result of poor drilling and blasting techniques, results in oversized, hard digging for the excavating equipment. The result is higher secondary blasting costs, reduced excavation efficiencies, higher excavation costs, higher road repair costs, and higher haulage equipment costs. The main objective of evaluating drill fleet requirements, therefore, is to determine the equipment needed to minimize the operational constraints, caused by blasthole drilling, on other facets of open pit mining at the lowest cash flow costs.

Jan 1, 1979

By D. E. Mikkola, G. E. Poquette

X-ray diffraction was used to study the growth of antiphase domains in quenched (or "disordered&apos;? Cu3Au annealed in the range 300" to 385°C. Measurements of the long-range order parameter indicate that a high degree of order is established rapidly at these annealing temperatures. Average domain sizes obtained by Fourier analysis of the superlattice peak profiles show that antiphase domain growth is analogous to classical metallurgical grain growth (i.e.,fol-lows a D2 US kt relation). The activation energy for the growth process, 44 kcal per mole, indicates diffusion control. The anisotropy in the observed effective domain sizes from seven different superlattice peaks can be best accounted for by a configuration involving antiphase domain boundaries on {100 } and (111 ) planes, where the (100) boundary is the low-energy type, {lOO} Type I. The disappearance of the (111) antiphase domain boundary during domain growth is dif-fusion-controlled with an activation energy of 48.5 kcal per mole. On the other hand, the behavior of the parameter representing the amount of (100) Type I boundary indicates that there may be a conversion of the higher-energy boundary types to (100) Type I boundary during the growth process. THE order-disorder transformation in the alloy Cu3Au has been the subject of numerous experimental investigations over the past 40 years, most of which have involved measurements of convenient physical properties such as resistivity and hardness as a function of heat treatment. The interpretation of these data is difficult because measurements of this type have the disadvantage that they yield total changes resulting from several different atomic processes which may be occurring simultaneously. The present study was undertaken to establish by means of X-ray diffraction the kinetic behavior of one of these atomic processes, namely antiphase domain growth. In the completely ordered state Cu3Au has an L12 structure which can be described as an fcc type arrangement of atoms in which the gold atoms occupy the corner sites and copper atoms the face-centered sites. Three other equivalent unit cells can be formed by allowing two gold atoms to occupy opposite face-centered sites and copper atoms to occupy the remaining corner and face-centered sites. These four different unit cells are crystallographically related by a shift of 1/2<110>. Quenching from above the critical temperature (392°C) to room temperature retains the "disordered" structure-a "random" arrangement of copper and gold atoms having the fcc structure. Annealing the disordered alloy below the critical temperature causes ordering to occur through the nucleation and growth of ordered regions at different points throughout the crystal. These small ordered regions, each based on one of the four types of unit cell, grow by consuming the disordered material until they impinge to form a domain structure in which the domains are antiphase by a shift of 4(110). The boundaries between domains are called antiphase domain boundaries (APDB) and they have been found to lie primarily on &apos; {100} planes in highly ordered Cu3Au. Two types of APDB can occur on ( 100) planes. An example of an (001)1/2[110] shear type APDB is shown in Fig. I. Because of the layer of copper atoms separating the domains in this case, there are no near-neighbor violations across the boundary and it has a low surface ener This type of APDB will be referred to as {lOO} Type I. Fig. 2 shows an (001): [011] climb type APDB, (100) Type 11. This type of APDB can be described as resulting from removal of a (200) plane followed by a shift of 1/2<110> to bring the separated regions together. Near-neighbor violations across (100) Type I1 APDB make it a high-energy boundary. There are numerous other possible APDB configurations; however, it is only worth noting that shear type APDB can be produced on the (111) plane by motion of a normal fcc dislocation with Burgers vector 1/2<110> through the ordered structure. A (111); [110] shear type APDB is shown in Fig. 3. Again, because of the near-neighbor violations this type of APDB has a higher energy than the (100) Type I APDB.

Jan 1, 1970

By D. Harrington

One of the main functions of the United States Bureau of Mines is to obtain and disseminate information that will promote safety in and around mines, and the health and safety of employees engaged in mining. Serious health and safety conditions arising out of contamination of air in metal mines by dust, fumes from explosives and fires, and absorption of oxygen by timber and other causes, led the Bureau of Mines to cooperate with the United States Public Health Service in the study of dust occurrence and ventilation in metal mines in various parts of the country. A. J. Lanza, of the U. S. Public Health Service, and Edwin Higgins, mining engineer of the U. S. Bureau of Mines, in 1915, made a detailed study in the mines around Joplin, Mo.; their reports1 are now available. In 1916, 1917, and 1918, a similar study was made in the Butte, Mont., region by Doctor Lanza and D. Harrington, mining engineer of the Bureau of Mines; the reports of this investigation will be available soon. In 1919, the study was continued in some of the Arizona mines by D. Harrington, of the Bureau of Mines, and G. E. McElroy and R. A. Koronski, of the U. S. Public Health Service, the underground work being completed in June, 1920; these reports are now being prepared for publication. In the spring of 1920, the scope of the work was extended, with D. Harrington in charge as supervising mining engineer, and short studies were made in the mines of the Coeur d&apos; Alenes in Idaho and around Oatman, Ariz. It is hoped soon to extend the work into mines of California, Nevada, Colorado, Michigan, and Minnesota, and ultimately to other states should the mining companies be willing to coöperate. While the coöperation of the mining companies is necessary, the companies are put to practically no expense other than supplying a guide to accompany the investigators underground, the length of time spent

Jan 1, 1922

By H. R. Hanley, C. Y. Clayton

One of the most important contributions to the art of electrolytic zinc production has been the aluminum cathode. This has been used in all major production since its commercial developmerit in 1916. The literature is replete with important contributions, some of which may have been replaced by better ones, but thus far there appears no other metal to take the place of aluminum as cathode material. Credit for this innovation goes to Sherard Cowper Coles, of England. When experimenting on the extraction of zinc from Broken Hill ore at Hayle, Cornwall, in 1893, he introduced the aluminum cathode, and in 1895 obtained . British Patent No. 20073. The reason for this dominant position of aluminum as cathode material in zinc sulphate electrolytes is its unique surface, to which electro-deposited metals usually do not tightly adhere. The metal must be of high purity for best results. It is possible to develop a starting-sheet method by which thin electro-deposited sheets of cathode zinc are made and stripped from a smooth aluminum plate. Such thin sheets of cathode zinc may be used as are the copper sheets in standard electrolytic copper-refinery practice. However, complications arise because of extreme warping of the zinc starting sheet after a few hours of deposition, and thus special arrangements must be made to compensate for this abnormality. Reasons advanced for the ease with which deposits may be stripped from the aluminum surface are: (1) presence of a permanent oxide coating on the metal, (2) failure of aluminum to alloy with the deposit, (3) differences in coefficient of expansion, (4) high position of aluminum in the electromotive series of metals. Of these reasons for nonadherence, which may have some application in general deposition, probably the oxide coating has the greatest bearing on the subject of zinc deposition. Adherence of zinc deposits has occasionally had serious effect on tonnage and production costs. Opinions have differed as to the specific

Jan 1, 1944

By H. L. Bilhartz, H. F. Dunlap, C. R. Bailey, Ellis Shuler

Data are presented which indicate that the saturation exponent, n, in the equation, R. = R100S-11, relating core resistivity, I:,. to the resistivity at 100 per cent saturation. R100. and to the saturation, S. may vary appreciably from the value of two which is usually assumed for this exponent when interpret ing well logs. Values ranging from one to two and one-half have been found on (.ore sample investigated to date. Attempts to correlate this saturation exponent with porosity or permeability of the core have not been successful. The saturation exponent is apparently not a function of the interfacial tension between the brine and the displacing fluid. Some evidence is given indicating that the resistance of the core is not a unique function of the saturation but depends upon the manner in which this saturation was achieved. Equipment and technique are discussed for measurement of resistivities in core plugs in which water saturation can be varied. lNTRODUCTION A number of investigations of the resistivity-saturation relationship for un-c~~nsolidated sands and consolidated (.ore samples have been reported in the literature. According to most of these: R. = R¹ººS², where R² = the resistivity of a formation at saturation S, and R¹ºº= the resistivity of the formation at 100 per cent water saturation. Much of this work was (lone on unconsolidated sands desaturated by gas or oil. Hen-clerson and Ynster worked exclusively with dynamic systems, flowing oil or gas through consolidated cores. There is some doubt as to how well this reproduces static reservoir conditions. Jakosky and Hopper³ onsidered also the case of consolidated core plugs, but the oil-water distribution in the emulsions which they used to saturate their cores is almost certainly different from that occurring in reservoirs. Recently Guyod quotes the results of some Russian work which indicates that n may vary from 1.7 to 4.3. No experimental details of this work are available. In connection with electric log interpretation it is important to know the value of the saturation exponent. For example, if in a given reservoir it is found that the resistivity is three time.; the resistivity observed when the reservoir is 100 pel. cent &apos;saturated with water, this fact would be interpreted as indicating a water saturation of 33 per cent if the saturation exponent were 1 and a water saturation of 6-1 per cent if the saturation exponent were 2.5. EXPERIMENTAL METHOD In the work to be described it was assumed that reservoir conditions are most nearly obtained when core plugs are desaturated by the capillary pressure technique referred to in numerous places in the literature, as for example. in Bruce and Welge&apos;s paper.&apos; In this technique the core. saturated 100 per cent with brine, is placed in contact with a ceramic disc permeable to brine but not to the displacing medium for the displacement pressures used. Pres-ure is then applied to the displacing medium and brine forced out of the core through the ceramic disc. Fig. 1 shows the core plug in place in the cell in which resistivity and saturation measurements are made. Fig. 2 shows the schematic electrical diagram wed to make resistivity measurements on the core plug. A four-electrode type circuit is used, employing a Hewlett-Packard model 400A. AC vacnum tube voltmeter. The 60-cycle AC current througli the core is adjusted to 1 milliampere and measured by noting the voltage drop across the calibrated 100-ohm resistor. The vo1tages appearing at probes 1, 2, 3, and 4 are then successively measured. Voltage drops across the top, center, and bottom portions of the core are obtained by sublracting the voltages appearing at successive probes. This technique avoids any polarization or other high contact resistance phenomena which may develop at the current input electrodes. Resistances which may develop between the core and the probes, and which are small compared to the 1-megoam input impedance 01&apos; the vacuum tube voltmeter will (obviously not affect the measurements allpreciably. Any very appreciable resistallces which may develop at any of the probe wires are detected and allowed for by inserting a 1-megohm resistor in series with the voltage measuring probe. If the probe resistance is actually zero, the new voltage measured after insertion of the I-megolim resistor will be approximately one-half of that previously measured. since the input impedance of the vacuum tube voltmeter is itself 1 megohm. If an! appreciable probe resistance has developed, the new voltage is found to be appreciably greater than one-half of the previously measured voltage. Such probe resistance; have been found to develop only occasionally and usually can be traced to poor connections betwern the core

Jan 1, 1949

By C. R. Bailey, H. F. Dunlap, Ellis Shuler, H. L. Bilhartz

Data are presented which indicate that the saturation exponent, n, in the equation, R. = R100S-11, relating core resistivity, I:,. to the resistivity at 100 per cent saturation. R100. and to the saturation, S. may vary appreciably from the value of two which is usually assumed for this exponent when interpret ing well logs. Values ranging from one to two and one-half have been found on (.ore sample investigated to date. Attempts to correlate this saturation exponent with porosity or permeability of the core have not been successful. The saturation exponent is apparently not a function of the interfacial tension between the brine and the displacing fluid. Some evidence is given indicating that the resistance of the core is not a unique function of the saturation but depends upon the manner in which this saturation was achieved. Equipment and technique are discussed for measurement of resistivities in core plugs in which water saturation can be varied. lNTRODUCTION A number of investigations of the resistivity-saturation relationship for un-c~~nsolidated sands and consolidated (.ore samples have been reported in the literature. According to most of these: R. = R¹ººS², where R² = the resistivity of a formation at saturation S, and R¹ºº= the resistivity of the formation at 100 per cent water saturation. Much of this work was (lone on unconsolidated sands desaturated by gas or oil. Hen-clerson and Ynster worked exclusively with dynamic systems, flowing oil or gas through consolidated cores. There is some doubt as to how well this reproduces static reservoir conditions. Jakosky and Hopper³ onsidered also the case of consolidated core plugs, but the oil-water distribution in the emulsions which they used to saturate their cores is almost certainly different from that occurring in reservoirs. Recently Guyod quotes the results of some Russian work which indicates that n may vary from 1.7 to 4.3. No experimental details of this work are available. In connection with electric log interpretation it is important to know the value of the saturation exponent. For example, if in a given reservoir it is found that the resistivity is three time.; the resistivity observed when the reservoir is 100 pel. cent &apos;saturated with water, this fact would be interpreted as indicating a water saturation of 33 per cent if the saturation exponent were 1 and a water saturation of 6-1 per cent if the saturation exponent were 2.5. EXPERIMENTAL METHOD In the work to be described it was assumed that reservoir conditions are most nearly obtained when core plugs are desaturated by the capillary pressure technique referred to in numerous places in the literature, as for example. in Bruce and Welge&apos;s paper.&apos; In this technique the core. saturated 100 per cent with brine, is placed in contact with a ceramic disc permeable to brine but not to the displacing medium for the displacement pressures used. Pres-ure is then applied to the displacing medium and brine forced out of the core through the ceramic disc. Fig. 1 shows the core plug in place in the cell in which resistivity and saturation measurements are made. Fig. 2 shows the schematic electrical diagram wed to make resistivity measurements on the core plug. A four-electrode type circuit is used, employing a Hewlett-Packard model 400A. AC vacnum tube voltmeter. The 60-cycle AC current througli the core is adjusted to 1 milliampere and measured by noting the voltage drop across the calibrated 100-ohm resistor. The vo1tages appearing at probes 1, 2, 3, and 4 are then successively measured. Voltage drops across the top, center, and bottom portions of the core are obtained by sublracting the voltages appearing at successive probes. This technique avoids any polarization or other high contact resistance phenomena which may develop at the current input electrodes. Resistances which may develop between the core and the probes, and which are small compared to the 1-megoam input impedance 01&apos; the vacuum tube voltmeter will (obviously not affect the measurements allpreciably. Any very appreciable resistallces which may develop at any of the probe wires are detected and allowed for by inserting a 1-megohm resistor in series with the voltage measuring probe. If the probe resistance is actually zero, the new voltage measured after insertion of the I-megolim resistor will be approximately one-half of that previously measured. since the input impedance of the vacuum tube voltmeter is itself 1 megohm. If an! appreciable probe resistance has developed, the new voltage is found to be appreciably greater than one-half of the previously measured voltage. Such probe resistance; have been found to develop only occasionally and usually can be traced to poor connections betwern the core

Jan 1, 1949

By Frank L. Gaddy

Falls of roof account for over 50% of the fatalities that occur in coal mines in the US. Thus, roof control is one of the more important phases of underground mining. In reality, the control of roof influences the system of mining and is a major determinant of the width and spacing of working places, operations at the mining face, ventilation control, and surface subsidence. Frequently, the control of roof is the largest single cost item. Roof control is a never-ending task, not only at the working face and throughout sections where men are working, but along haulways and air- ways that must be maintained for the life of the mine. This is because roof, even the best of roof, slowly deteriorates, so it must be frequently examined, with corrective support applied when needed. As coal is excavated, stresses are set up in the roof because the previous equilibrium is upset, resulting in pressures that cause fractures and slight movements that are frequently hard to detect. Unless the immediate roof, in the excavated area, is given support by artificial means, it might fall or there might be a succession of falls, varying from a thin scale to several feet, depending on the nature of the top. Exceptions to this are those rare mines, or sections of mines, with hard, strong, nonweathering roof that requires no support. GENERAL CLASSIFICATION OF ROOF There are two broad types of roof as far as support is concerned: the immediate roof above the coal and the main roof. Artificial support for mining purposes is only concerned with the immediate roof as nothing except large blocks of solid coal, or massive concrete, will support the main roof. The immediate roof is generally a few feet thick but can vary from inches to 6.1 m (20 ft) or more. There are examples of where there is no immediate roof as the massive main roof lies directly on the coal

Jan 1, 1981

By Thomas T. Read

REVIEWING the events of the year in mineral industry education, a certain amount of either amusement or irritation, depending upon one&apos;s viewpoint, can be derived front the section dealing with mining engineering in Walter B. Pitkin&apos;s "New Careers for Youth" which appeared during the summer. On p. 102 Prof. Pitkin says "The mining engineer is disappearing .. . because . . . the profits of mining decline rapidly with the age of the mine." In his "Are Too Many Students Taking Mining Courses?" published in the May issue of this journal, W. B. Plank presents statistics showing that there was some decline in student enrollment in mining and metallurgical courses in 1932-3 and 1933-4, compared with 1930-31, but that the decline was proportionate to that in all lines of engineering study combined. Informal reports from Institute officials that have visited tinning schools during the fall indicate that nearly all schools showed an increase in enrollment for the current school year. Pre else statistics will be made available by Prof. Plank at the Annual Meeting. His study last year provided, for the first time, accurate and complete information on the enrollment in mining schools, as well as the number of various courses and the degree granted, and it will be interesting to have comparable figures for this year. Prof. Pitkin&apos;s views on the future opportunities for mining engineers were discussed at some length - on pages 439-4.40 of the November issue of MINING AND METALLURGY.

Jan 1, 1935

By John Randall

IT is a fair assumption that the main purpose of any diagram is to present facts to the eye in more convenient form than they could be tabulated in figures, and this implies that a screen diagram should set forth in a convenient and familiar form certain facts as to the character of the material screened regardless. of the screen standards used, and it seems particularly desirable on account of the well-known differences in the size of openings of different laboratory screens having the same nominal mesh. In making screen-sizing tests it has been customary to first consider the largest size, going progressively to the smallest, and this seems to have led to the practice of placing the smallest size at the right-hand side of the diagram and .measuring toward the left. This is not altogether convenient and might he changed to advantage, but for illustration I have preserved that feature and present the method just as I have used it for over 2 years. Direct plotting in which the intervals on the diagram are proportional to the sizes of the openings, or to the arithmetical difference between them, seems at first sight more natural and is extremely useful for certain purposes, but it makes the diagram very crowded in the smaller screen sizes while the larger ones occupy unnecessary space, snaking it less convenient for most purposes, and for that reason the method here discussed is the well-known one based oil the ratio between openings, and is plotted in logarithms instead of the direct values of the screen opening. Metric coordinate paper is convenient, although .not necessary; it is recommended, as the use of the metric scale gives a diagram of convenient size, but any scale of equal. parts may be used. In order to make any screen diagram translatable in terms of any standard screen size, it is necessary to have rioted on the size scale of the diagram all points corresponding to the openings of the various screens in question, instead of a single series &apos;of points as with a single screen. In drawing the diagram assume 0.001 in. (corresponding approximately to 0.025 mm.) to be the smallest size to be considered and fix this as the zero point at the lower right-hand corner of the diagram. This value is selected because the values are to be plotted in logarithms, and the logarithm of 1 is zero. Then lay off to the left for the position of each screen

Jan 9, 1917

By E. L. Jr. Spain

This study was undertaken primarily to determine the reasons for certain varia-tions in the soundness of gravel aggregate taken from a number of widely separated points on the Tennessee River. Under laboratory tests, concrete made with these various aggregates it some cases exhibited marked differences in specific gravity, crushing strength and durability. Gravel was studied in detail from Chattanooga to the mouth of the river and to a lesser extent from Chattanooga to Knoxville. The causes of variation were determined, and variational curves were constructed, which accurately predict the composition of aggregate at any point on the river and therefore its relative suitability for use in concrete. The aggregate at any point is suited for making sound concrete; the variance of suitability is due to varying percentages of sandstone and weathered material, which are less sound than other constituents. INTRODUCTION When it is stated that the aggregate from one point is inferior to that from another, it must not be construed to mean that the gravel is unfit for use. The classification upon which such aggregate ratings are based is largely arbitrary, thus inferiority or superiority is merely relative. Samples of gravel were selected from stockpiles at the dams of the Tennessee Valley Authority under construction, from the stockpiles of private companies along the river, and from samples taken by the gravel-exploration barge of the T. V. A. One hundred or more gravels of each size studied were selected at random by taking small handfulls while walking around the piles. These were separated according to the differ-ent mineral or rock constituents present, and the percentage of each constituent was obtained by counting. At least three such samples were taken from each size, and all checked with one another within a few per cent. The results were then averaged. In order to check on the accu-racy of such sampling, a whole sample was taken from one point after the gravel was thoroughly mixed. The two methods gave approximately identical results.

Jan 1, 1937

By Frederick H. McDowell

Notwithstanding the rapidly increasing use of copper, due to the extension of its applications within the last few years, the fact of its continued steady decline in price stands prominently forward. Nor have we far to look for the cause. The opening up of the vast copper-deposits of our Western States and territories is the metallurgical event of the last five years. We must not forget, in looking at the new field, the changed conditions under which the metallurgist must work, as compared with the previous history of copper-smelting. Until quite recently, the copper-plants of the country were few in number; and the concentration of industry implied in this circumstance, allowed many practices which must be considered as inapplicable to the conditions offered at the many mines which have of late become such important factors in our copper-production. Thus the old style of smelting the ore for matte in brick furnaces, involving great expense and time in their construction, while deservedly losing repute even in its old strongholds, in comparison with water-jacketed smelters, is still in some cases adhered to. But to put up such furnaces at Western mines would be out of the question. The large and continual need of brick for relining and repairs acts as an effectual bar, when such furnaces come into competition with the light, portable, and easily managed water-jacketed smelters of improved types. It is owing to the present unusual activity, manifested in the opening up of new copper-properties and the erection of plants for working them, that any improvement in furnace-construction is of especial interest and importance. The method in which processes are copied without variation in the rapid growth of metallurgical interests, has already been strikingly exemplified in the history of

Jan 1, 1885

By W. C. Lilliendahl, H. W. Highriter

The method developed and used in this laboratory for the production of metallic uranium of such purity that it is ductile and can be cold-worked to fine wire or thin sheet by rolling has already been described1. It consists, briefly, in the electrolysis of potassium uranous fluoride (KUF5) in a molten bath composed of equal parts of sodium and calcium chlorides. The bath is contained in a graphite crucible, electrically heated to about 775° C., the crucible serving as the anode for electrolysis. The uranium deposits in finely divided form mixed with salts upon a molybdenum cathode suspended axially in the crucible. Upon completion of the electrolysis the cathode is removed and cooled, the deposit broken and leached to remove salts. The metal powder is washed with water and dilute acetic acid and dried in vacuo to reduce oxidation. Suitable amounts are then prcssed into pellets and heated in vacuo by high-frequency induction, using a thoria crucible. The uranium melts and flows away from the pellet, leaving a shell, which is largely oxide. In general, this method produced a satisfactory metal, although at times tile uranium was not ductile. The investigations recorded here were made to find the cause of this embrittlement. It was known that the metal generally contained carbon, derived from disintegration of the graphitc crucible in which electrolysis is conducted, in amounts dependent upon the thoroughness of washing the metal powder. Accordingly, analyses of several samples, both ductile and brittle, were made by combustion and absorption of CO² in Ascarite, using air instead of oxygen to reduce the temperaturc and rate of oxidation. Several of these metals were made with special precautions to reduce carbon content. The results show no relationship between ductility and carbon content in the range investigated (Table 1). Microscopic examination of ductile and brittle metals disclosed the presence in the latter of an intergranular constituent, harder than the matrix and less easily oxidizable, which was absent from the ductile material. Rough estimates of the volume of the hard constituent

Jan 1, 1935

By W. C. Lilliendahl, H. W. Highriter

The method developed and used in this laboratory for the production of metallic uranium of such purity that it is ductile and can be cold-worked to fine wire or thin sheet by rolling has already been described1. It consists, briefly, in the electrolysis of potassium uranous fluoride (KUF5) in a molten bath composed of equal parts of sodium and calcium chlorides. The bath is contained in a graphite crucible, electrically heated to about 775° C., the crucible serving as the anode for electrolysis. The uranium deposits in finely divided form mixed with salts upon a molybdenum cathode suspended axially in the crucible. Upon completion of the electrolysis the cathode is removed and cooled, the deposit broken and leached to remove salts. The metal powder is washed with water and dilute acetic acid and dried in vacuo to reduce oxidation. Suitable amounts are then prcssed into pellets and heated in vacuo by high-frequency induction, using a thoria crucible. The uranium melts and flows away from the pellet, leaving a shell, which is largely oxide. In general, this method produced a satisfactory metal, although at times tile uranium was not ductile. The investigations recorded here were made to find the cause of this embrittlement. It was known that the metal generally contained carbon, derived from disintegration of the graphitc crucible in which electrolysis is conducted, in amounts dependent upon the thoroughness of washing the metal powder. Accordingly, analyses of several samples, both ductile and brittle, were made by combustion and absorption of CO² in Ascarite, using air instead of oxygen to reduce the temperaturc and rate of oxidation. Several of these metals were made with special precautions to reduce carbon content. The results show no relationship between ductility and carbon content in the range investigated (Table 1). Microscopic examination of ductile and brittle metals disclosed the presence in the latter of an intergranular constituent, harder than the matrix and less easily oxidizable, which was absent from the ductile material. Rough estimates of the volume of the hard constituent

Jan 1, 1935

By W. D. Robertson

The effect of a tensile stress in accelerating the corrosion-cracking of certain alloys of aluminum, magnesium and iron is widely recognized. The literature is extensive and it is only necessary to cite the recent ASTM—AIME Symposium,l There are, however, no experimental data conclusively showing the effect of an applied compres-sive stress. It has been inferred, from the generally accepted mechanism postulated for the acceleration of corrosion by an applied tensile stress, that a compressive stress would inhibit corrosion. Furthermore, it is observed that corrosion cracks invariably start on the tension side of a specimen bent as a simple beam. This observation has led to the suggestion that shot-peening of the surface would inhibit or eliminate stress-corrosion cracking in magnesium alloys and aluminum alloys. However, the degree of inhibition produced by shot-peening does not appear in the literature except for the above qualitative references which are unsupported by experimental data. Moreover, while peening undoubtedly imposes a surface compressive stress, the alteration of the surface by the severe cold-working involved complicates the interpretation of the results with respect to the effect of a compressive stress alone. If corrosion proceeds along preferential paths produced by precipitation from solid solution, it is easily understood how a component of tensile stress normal to these paths will accelerate the corrosion and lead to ultimate failure. On the other hand, if the metallurgical conditions are such as to create preferential corrosion paths, then it is difficult to understand how a compressive Stress, produced by shot-~eening Or any other method, can inhibit corrosion except in the sense that the accelerating factor is removed. Thus, if the basic corrosion mechanism remains unaltered, there should be little or no inhibition resulting from the application of a compressive stress if inhibition is defined as the reduction of the degree of corrosion below that experienced in the unstressed state. The experiment described below was performed to isolate and evaluate the effect of a compressive stress on the corrosion and stress-corrosion cracking of a high strength aluminum alloy (24-S); and, by a Comparison of the corrosion rate of &apos;pecimens subjected to an applied tensile stress, compressive stress and in the unstressed state, to clarify the role of stress in intergranular corrosion. Experimental Procedure The experiment was conducted with the aluminum-copper-magnesium alloy known as Duralumin 24-S.* Specimens, 0.064-in. thick, in the form of I-in. wide strips sheared parallel to the rolling direction,

Jan 1, 1949

By M. N. Short, I. A. Ettlinger

M. N. Short,? Washington, D. C., and I. A. Ettlinger, New York&apos;, N. Y. (New York Meeting. February, 1926) The Pioneer mining district, better known as the Superior district, from its principal town, is located in Pinal County in south-central Arizona about 80 miles east of Phoenix and 22 miles west of Miami. The region embraces a, confused series of mountain ranges of moderate height but precipitous slopes, between which lie irregular valleys filled with alluvial detritus. The average elevation is about 2800 ft. in the valleys and the summits of the ranges reach 2000 ft. above this. Superior is situated on the north bank of Queen Creek where it emerges from one of those ranges into a more or less basin-shaped valley. The drainage of Queen Creek is to the southwest and the flood waters spread out and lose themselves in the desert northeast of Florence, the county seat of Pinal County. The Magma copper mine (Fig. 1) is the only producer in the district. Although discovered in 1875, its production has been important only

Jan 1, 1927

By W. A. Rachinger, C. Graeme-Barber, R. L. Bell, R. C. Gifkins, T. G. Langdon

R. L. Bell. C. Graeme-Barber, and T. G. Langdon (Imperial College. London)— The internal-marker technique developed by Ishida, Mullendore, and Grant has enabled them to make some interesting observations on the role of grain boundary sliding in creep. Certain of their conclusions are supported by our own similar experiments on Magnox AL80 (Mg-0.8 pct Al) specimens containing several markers, full details of which will be published shortly. In particular, we would agree that: 1) Using the internal-marker method, there is little difference between Egb, the contribution made by grain boundary sliding to the total strain, measured at the surface and in the interior of a specimen. 2) The measurement of the change in grain shape during creep leads to an underestimation of the grain strain, and hence an overestimate of Egb.. However, the discrepancies between the values obtained by this method and those obtained using markers are not as great as Tables IV and V of Ishida, Mullendore, and Grant would suggest; see below. The authors&apos; Eq. [I] which is used to compute Egb from the transverse displacement ITt of longitudinal marker lines is not given in the reference quoted,&apos;&apos; where in fact only relations for Egb in terms of the longitudinal displacements of markers were developed. As first pointed out by Gifkins and Gittins,12 a problem arises in the use of longitudinal markers. and we will show that Eq. [1] underestimates Egb in the interior by a factor of 2. In Fig. 12, two grains X and Y have become displaced by the vector AC which lies in the plane of the grain boundary ABCD separating them. A longitudinal marker has been split by this displacement into the two parts LA and CAW. If these were surface grains the transverse displacement, Ut. could be obtained by viewing normal to the "top" surface. If interior, a section through CD parallel to the top surface would show the same result, provided the longitudinal internal markers delineated planes normal to the top surface. (This is an essential feature of the markers not pointed out in the paper.) The important point is that the quantity Ut tan B (=BE=DF) only gives the elongation due to the one component AB of the sliding. The elongation V/tan 8 (=GD) due to the component AD must be added to produce the total elongation resulting from the boundary displacement AC. If the grain boundaries in a poly crystalline specimen are randomly oriented with respect to the stress axis it follows that In the interior of a specimen the two sums depend for any difference simply on the definitions "top&apos;&apos; and "side". On the surface it is possible that the two sums could differ somewhat. In either event it is clear that the authors&apos; Eq. [I] underestimates Egb, and in the interior, and perhaps at the surface also, the appropriate formula is given by The suggested modification probably makes little difference to the comparison between surface and interior values of Egb where both were obtained from marker measurements. For the estimation of the correction becomes extremely important where the boundary contribution is large. In our own work on magnox, for example, we have obtained values of as large as 70 to 80 pct, under conditions of high temperature and low strain rate. The use of planar markers transverse to the stress axis would simplify the comparison between surface and interior behavior, because measurement of the longitudinal offsets Us of these markers could be summed to give the total elongation di-

Jan 1, 1965

By F. L. Kadey

The petrographic microscope is a well known tool in nonmetallic materials research. Its utility in the study of perlite is, therefore, not surprising. Valuable information has been derived on samples smaller than those needed by most other methods of analysis. Relatively little has been written about the petrographic analysis of perlite for two chief reasons: perlite is a material of relatively new commercial interest; and the independent prospector or expander, responsible for most studies, attacks the problem from an immediately practical viewpoint. Hence, petrographic studies of perlite have been limited to academic aspects. Rapid development of commercial interest in perlite has changed the situation. Mining and expanding companies now have effective research tools; and such instruments as the petrographic microscope have added materially to practical knowledge about this mineral. Petrographic techniques employed in the search for suitable deposits screen worthless samples rapidly and inexpensively compared with large-scale expanding tests. Hence, only samples of potential quality are selected for further testing. Field appearance of perlite reveals little about its true identity and practically nothing about its commercial significance. Consequently, any method that can be used to separate the "wheat from the chaff" is valuable. In addition to the conventional thin-section and immersion methods used in evaluation of crude perlite, petrographic microscope techniques have contributed to the study of expanded perlite. When apparently similar materials behave as differently in use as perlites, the internal structure of the individual particles becomes a subject of great interest. The purpose of this paper is to discuss some of the petrographic techniques used to characterize this material. WHAT IS PERLITE AND WHERE IS IT FOUND? Perlite was at first described as a glassy rhyolite possessing a pearl-like luster and exhibiting numerous concentric cracks resembling an onion skin appearance.&apos; Fig. 1 is an illustration of perlite in this classical sense. Recently, its definition has taken on a commercial connotation: an acid volcanic containing 2 to 5% of combined water that, when heated rapidly to the softening point of the glass, coincident with the volatilization of the combined water, expands or "pops". Fig. 2, although hardly fitting the classical description given by Johannsen, is an excellent perlite in the commercial sense. It may occur as a breccia in rhyolite, (Fig. 3) and sometimes it is associated with obsidian nodules, as illustrated in Fig. 4. The term perlite includes the expanded product also. In the United States, perlitic rocks are restricted to the western part of the country. 2 Elsewhere in the northern hemisphere, European occurrences follow a belt beginning in Iceland and extending in a southeasterly direction through Scotland, County Antrim in Northern Ireland, and through the Massif Central of France, Germany, and Hungary; thence, into Italy and Sardinia, and through some of the Aegean Islands. A general oriental belt extends from Japan through Australia and New Zealand. Rhyolitic areas elsewhere in the world undoubtedly contain perlite.

Jan 1, 1963