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By H. L. Hopkins, L. W. Kempf

Aluminum alloys, in common with most other metallic alloys exhibit slight density changes with variations in temper achieved by heat-treatment, which usually are the result of the variation with temperature of the solid solubility of the various alloy constituents in the solvent metal. Changes in density are, of course, most marked in conditions that exhibit the maximum variation in solid solubility of the constituents. In aluminum alloys these conditions usually prevail when alloys are quenched from just below the melting point and are subsequently reheated to some lower temperature. Since commercial heat-treatments may be of this type, the density changes, although small (of the order of 0.3 per cent which may be compared with the density change of about 1.5 per cent involved in the transformation, at ordinary temperatures, of gamma to alpha iron during the hardening of steel), may in some specific applications necessitate quantitative consideration. This phenomenon has been investigated extensively in the laboratories of the Aluminum Company of America over a period of years. A number of investigators '-5t have observed and measured the density changes accompanying the heat-treatment of aluminum alloys. In general, these changes have not been directly translated to dimensional changes and their significance has been overlooked to some extent by the practical user of aluminum alloys. It is the purpose of this paper to consider some practical aspects of these density and dimensional changes. These density changes were given some consideration in 1917 in connection with the use of aluminum-alloy pistons for aircraft engines. Investigations by Lea and Shakespear were reported in some confidential reports of the British Advisory Committee for Aeronautics, Light Alloys Subcommittee, during 1917 and 1918, which are not generally available in published form. In these reports the term "permanent growth" was applied to the dimensional changes occurring in the pistons at the tem-

Jan 1, 1936

By J. Lemish

Considerable research has been conducted in Iowa in recent years on carbonate aggregate reactions in concrete. Groups outside the state have become active in the past few years after recognizing that certain carbonate aggregates are reactive in concrete. The paper reviews the nature and extent of the problem and a summary is given of research by the author and his associates. In recent years considerable research has been conducted in Iowa on carbonate aggregate reactions in concrete. The research is part of a general study on the behavior of carbonate rocks as concrete aggregates sponsored by the Iowa State Highway Commission. Research on carbonate aggregates has not been extensive in the past. Laughlin1 published one of the first papers on carbonate aggregates in 1928 and was followed by others, notably H. S. sweet,' B. Mather et al.,3 and K. Mather4 who emphasized the petrographic and physical aspects of carbonate aggregates. The recognition that certain carbonate aggregates are reactive in concrete has led to increased participation in research in the past few years by groups outside of Iowa. Significant contributions have been made by Swenson and Gillot,5 Hadley,6 and Chaiken and Halstead7 in work which emphasizes the expansion of carbonate rocks in alkaline environments and the reaction of carbonate minerals or dedolomitization with alkalies. It is the purpose of this paper to review briefly the nature and extent of the problem and present a summary of the research by the author and his associates. EXTENT AND NATURE OF THE PROBLEM The problem is the deleterious behavior of certain carbonate aggregates in concrete. It was first recognized in Iowa over 20 years ago when poor service records of certain highways were correlated with aggregates produced from the Glory and LeGrand quarries. Deleterious behavior of carbonate aggregates have been reported in Canada and, more recently, in New York and Virginia. As more knowledge is gained concerning the problem, distress in concrete related to carbonate aggregates may be recognized over a widespread area. The problem may potentially exist in the Midwest, the Northeast, and other parts of the country underlain by thick sections of carbonate rocks readily available for utilization as aggregates. The nature of distress in Iowa highways is a general weakening and deterioration of concrete characterized by progressive cracking and spalling away from joints and edges (Fig. 1). The distress generally occurs after 15 years of service and becomes progressively worse. Specimens of such concrete lack the toughness of sound concrete and break easily under the hammer in the manner of a soft brick. The matrix is generally highly carbonated and dark borders, or reaction shells or rims outline the argillaceous carbonate aggregate fragments (Fig. 2). Cracks observed in cores cut across the aggregates and concrete matrix. Although localized volume change (expansion) is associated with spalling, there is no definite field evidence in Iowa that excessive growth of concrete slabs has occurred. The experience with carbonate aggregates in Iowa indicates that the concrete deterioration is not the same as that induced by the alkali-aggregate reaction associated with siliceous aggregates. The principal difference is that in Iowa the distressed concrete is weakened throughout the slab. In concrete affected by the alkali aggregate reaction, the

Jan 1, 1961

By Wm. Hamilton Merritt

With the exception of the Vancouver Island coal, all the western coal-fields owe their present development to the completion, in the autumn of 1886, of the Canadian Pacific railroad. While it could not be expected that a very great deal could be accomplished in three years, enough has been done to pretty thoroughly establish the coal-bearing areas and their correspondence with those which have been developed to the south of the boundary along the lines of the transatlantic railroads in the United States. This summer I visited some of the important developments in the coal areas of Washington Territory, largely with the object of being better able to appreciate the corresponding coal-bearing areas in British Columbia to the north. In Western Canada, coal-bearing rocks have been found in three zones: 1. In the plains to the east of the Rocky Mountains and in the eastern flanking ranges, the coal occurs in the Cretaceous formation (including the Laramie). 2. In the interior plateau of British Columbia, the coal is found in the Tertiary formation. 3. On the coast of British Columbia, Cretaceous and Tertiary rocks are found carrying coal, and on the Island of Vancouver the well known Nanaimo coal has been worked for years in the first named formation. In all of these zones, the coals vary from lignites up to higher grades, the factor determining quality being the amount of pressute to which they have been subjected. The intensity of this pressure is generally shown by the disturbance which the coals exhibit, and, in many cases, is almost directly in proportion to the distance of the deposits from mountain ranges. This seems to be also the opinion expressed by Mr. Bailey Willis in connection with his Census Report on the coals of Washington. It has been elsewhere stated

Jan 1, 1890

By Kenneth B. Barnes, George H. Fancher

With the advent of water-flooding into active commercial usage in half a dozen areas in the Mid-Continent, the process passes the experimental stage and joins other methods and processes now in use which have been found suitable in one oil field or another in securing a more profitable production than has existed heretofore. The use of the air-gas lift, acidization, pressure maintenance and repressuring have presented their problems in economics and engineering during their introduction and extended application. When a new field is discovered or attention is turned to an old producing area in which these methods have not been considered, it is to be expected that these problems again arise. While some information is available as to the results obtained by water-flooding in a few localities in the Mid-Continent, the extrapolation of the sand data on which the results are based, for even generalized use in other areas and fields in the region, is hazardous. It is precarious to assume that the same physical condition will obtain in a sand over more than small, special areas. This paper is concerned with a brief presentation of such data as are available to the authors on existing operations, a survey of the region and extent to which application of the method seems most likely to be made, an estimation of the crude reserve existing for such operations, suggestions as to the importance of core analysis in guarding against uncontrolled floods and its usefulness in predicting ultimate recoveries, and comment on some of the economic phases involved. These subjects, as concerned with water-flooding in the East (notably Bradford, Pennsylvania), have been discussed in detail in the literature and are generally well known. Especially is this true of discussions of the mechanical equipment and various lease layouts and development patterns for water-flooding.

Jan 1, 1936

By Kenneth B. Barnes, George H. Fancher

With the advent of water-flooding into active commercial usage in half a dozen areas in the Mid-Continent, the process passes the experimental stage and joins other methods and processes now in use which have been found suitable in one oil field or another in securing a more profitable production than has existed heretofore. The use of the air-gas lift, acidization, pressure maintenance and repressuring have presented their problems in economics and engineering during their introduction and extended application. When a new field is discovered or attention is turned to an old producing area in which these methods have not been considered, it is to be expected that these problems again arise. While some information is available as to the results obtained by water-flooding in a few localities in the Mid-Continent, the extrapolation of the sand data on which the results are based, for even generalized use in other areas and fields in the region, is hazardous. It is precarious to assume that the same physical condition will obtain in a sand over more than small, special areas. This paper is concerned with a brief presentation of such data as are available to the authors on existing operations, a survey of the region and extent to which application of the method seems most likely to be made, an estimation of the crude reserve existing for such operations, suggestions as to the importance of core analysis in guarding against uncontrolled floods and its usefulness in predicting ultimate recoveries, and comment on some of the economic phases involved. These subjects, as concerned with water-flooding in the East (notably Bradford, Pennsylvania), have been discussed in detail in the literature and are generally well known. Especially is this true of discussions of the mechanical equipment and various lease layouts and development patterns for water-flooding.

Jan 1, 1936

By J. David Lowell

For his contributions to the field of mining geology, a striking example being his brilliant inte.rpretation of the complex geologic history of the San Manuel District (Pinal County, Arizona) which resulted in the discovery of the Kalamazoo orebody, and for his lecture, "Copper Resources in 1970." On a low desert ridge six miles south of Cairo lie the remains of the world's earliest known mining town, Maadi of predynastic Egypt. Copper artifacts, including ingots and an ax head which was spoiled in casting, indicate that copper ore was smelted in Maadi, and radiocarbon dating has placed the activity at least as early as 3300 B.C. This is the oldest documented date for the use of smelted copper, the first industrial metal used by man. Copper was mined and smelted almost as early in Asia Minor and on the island of Cyprus, but the greatest development of the budding copper industry was in Egypt, where specific historical records show that about 10,000 tons of copper metal were produced in the period from 1850-1580 B.C. Contrast this production of 35 tons per year with the present world production of about 7 million tpy. Evolution of the Copper Industry The copper ore mined by the ancients was high grade, and the copper produced was expensive. The copper industry began with narrow high-grade veins, but has evolved to the point where massive earth-moving equipment can be operated in huge, low-grade open pits. Indian pits dug before 3000 B.C. in the Lake Superior district were in ore which averaged 15% Cu. His- torical research done by Arthur Brant has shown that copper cost in the Middle East in 1840 B.C. in terms of 1970 labor dollars was about $25 per lb. The copper mined in northern Europe in 1540 in the time of Georgius Agricola averaged about 8% Cu, and the cost in 1970 labor dollars was about $10 per lb. Average ore grade had dropped to about 6% Cu in 1890, and the cost of copper in the U.S. in 1970 dollars was about 45¢ per lb. Butte, Mont., was a good example of the efficiently operated vein mines of this time. In 1906, Daniel Jackling made a dramatic entrance into the copper scene with the Bingham Canyon open-pit copper mine and proved that more profit was possible mining 2% porphyry copper ore on a

Jan 1, 1971

By C. G. McLachlan

ONE phase of the treatment in sulphide flotation, covered generally in a review of pulp pretreat-ment by S. A. Falconer,' is the matter of grinding mills as conditioners, a subject on which further study could be advantageous. Flotation with few exceptions is preceded by grinding, and the grinding and flotation capacity required for the treatment of an ore is based generally on laboratory or pilot mill tests. The results of such tests, it then is assumed, will be duplicated in practice, and a flowsheet is prepared on that basis. This assumption may be too broad, because in several instances laboratory flotation tests made on massive sulphide copper ores could not be duplicated in plant operation without the introduction of an aeration step, either in closed circuit with the grind-inging mills, or between grinding and flotation.~' " The differences between laboratory and plant grinding practice are considerable, and may be summarized as follows: 1—In the laboratory a relatively fine charge is ground in a batch mill. In the plant, the feed, usually considerably coarser than the ore tested in the laboratory, enters the grinding mill continuously and more often than not, oversize material in the mill discharge is retained in the grinding circuit by a classifier until the required degree of reduction has been effected. 2—The rate of grinding in the laboratory is slower than in plant practice. 3—The percentage of volume occupied by the pulp in a laboratory mill rarely approaches that of a continuous overflow type mill. This does not apply if the comparison is made with grate type mills, but in that case an even wider discrepancy between the relative rates of laboratory and mill grinding almost invariably exists. The present trend in grinding is toward larger mills to reduce the number of operating units. This accentuates the foregoing differences. It is believed that the trend toward larger mills is sound, but that when large mills are used, metallurgists should check their flotation circuits carefully to determine whether laboratory tests are being reproduced in

Jan 1, 1952

By J. Fageant, W. C. Ellis

All major regions in a progressively solidified germanium ingot were related through successive orders of octahedral twinning. The occurrence of lineage structure and the generation and survival of orientations are discussed. WHEN metals solidify in a temperature gradient a preferred orientation, or casting texture, usually develops. For example, in face-centered and body-centered cubic metals, most of the solidifying crystals have a cube axis, <loo>, aligned approximately in the freezing direction.&apos; In the solidification of diamond cubic metals this is not the case, since there is a natural tendency to form twin orientations.&apos; . In this paper, twinning relationships in cast germanium are described; some conclusions with respect to nucleation and survival of new orientations are discussed. In the first experiment, an ingot of high purity germanium, % in. in diameter and 1 % in. in length, was formed by progressive solidification4 from the bottom. The ingot was cut into 21 slices, each circular and about 0.030 in. thick. The top surface of each slice was polished and etched to reveal the regions of different orientations. Orientations were determined by the Laue X-ray back-reflection method. The uncertainty in the determined orientation relationships was l" or 2" when separate slices and films were involved. Frequently, however, advantage was taken of illuminating across a boundary when the patterns of the two orientations were recorded on the same film. The uncertainty then was less than 1". Often the twin relation was evident from common zones or individual reflections. Twin relationships were recognized through comparison of the values of the nine angles between cube poles of the two orientations with those calculated for successive orders of octahedral twinning. The birth and survival of each region can be reconstructed by visual examination of the etched slices shown in Fig. 1. In slice 21, where solidification began, many regions are present; these regions were found to be twin-related. Solidification appears to have started from one nucleus formed probably at the wall of the crucible.* Subsequent ori- entations were developed by nucleation on existing solid surfaces, and only in the twin relationship. This is an example of a general process of nucleation on a surface of existing solid and may be descriptively termed oriented nuc1eation.t In germanium, and other diamond cubic substances, the orienting habit is that of twinning on an octahedral plane.&apos; " The prevalence of straight traces in the polished surfaces of succeeding slices (Fig. 1) suggests that the twinning observed in the first slice persisted throughout solidification. This was confirmed by X-ray determinations. The schematic diagram of Fig. 2 shows the twinning relationships of the major regions throughout the ingot. Since all observed relationships are describable as twinning, it is reasonable to conclude that no random nucleation occurred in the freezing of the ingot. An alternative possibility would be that nucleation of a random nature occurred, but only those nuclei in twin orientations survived in subsequent growth. This is unlikely for there would be as good a reason for the growth of randomly oriented nuclei as for the growth of the many different orientations present after several orders of twinning. The many twin components contain a wide distribution of orientations, and did grow. In fact, they are the only ones found in the solid. The conditions of solidification of this ingot were such as to discourage nucleation of crystals wholly in the liquid. The rate of solidification was 0.125 in. per min. This slow rate of heat removal afforded but small opportunity for supercooling within the liquid —a condition needed to provide enough decrease in free energy for such nucleation. When nucleation occurs on an existing solid surface, the interfacial energy between the existing and forming regions is important in determining the new stable orientation. The new orientation is expected to be one which minimizes this energy. In the diamond cubic structure, the interface for the

Jan 1, 1955

By H. W. Mossman

Three aspects of magnetite smelting are discussed. The first is the working out of equilibrium conditions for eliminating sulfur. The second is the influence of magnetite solubility on the difficulty of tapping the reverb matte. The third is an approximation of the equilibrium conditions in the reverb gases which govern whether magnetite is mode or reduced in the reverb slag by these gases and by any iron sulfide in the slag. MAGNETITE has had a varied history in the Hurley Smelter since its start in 1939. Magnetite determinations on the smelter products are made regularly only on the monthly composite samples. Variations on the monthly averages are shown in Table I. Magnetite which drops from the slag and matte in the reverb has some slight bottom buildup which comes and goes, but no substantial accumulation from this source has been found at the end of a normal nine months&apos; furnace campaign. However, there has been some low grade magnetite bearing material mixed with considerable A1,0,, which has slid down from the bottom of the sloping flue between the reverberatory furnace and the waste heat boilers. This accretion has required drilling and blasting near the skimming end of the furnace. The magnetite has interfered with tapping at times. When the smelter was first started, tapping trouble from magnetite was extremely severe. Increasing the reverberatory furnace temperature by putting in an air preheater and a Dutch oven has helped greatly, although there still is occasional tapping trouble. When the present series of physical chemistry articles on copper smelting started coming out in 1950, they were read with interest, but no immediate application was seen for them. Results of some laboratory work in 1952 aroused a much stronger interest in this physical chemistry. A series of melts was made on some converter slags, which had magnetite in very large grain sizes, with the object of reducing the grain sizes in the slag, as it was known that it was easier to handle in the reverb in that condition. Anything done in the tests greatly reduced the grain sizes—even in the controls, where nothing was done except melt the slag and cast it. There was more magnetite in the slags after the tests than before, and with wide variations. There were no obvious reasons lor much of what happened in these tests. Much of the base material published in English in this field was made available for study. Recalculations were made on many of the type problems, and part of the data was reduced to local temperatures and compositions. Explanations were found for what happened in the 1952 series of tests on converter slags, and the same principles turned out to be a description of much of what magnetite does in the reverb. This article is to present the results of that study, from the viewpoint of applying the technical material in definite numerical form to the operating conditions in both the converters and the reverberatory furnaces at the Hurley smelter. Table I. Magnetite Variations on Monthly Averages, 1939 to 1955 Pet Magnetite Lowest Highest Average Converter slag 13.6 43.3 25.4 Roverb slaa 2.7 20.9 8.7 Matte 28 15.9 98 In general it was found that magnetite is made or reduced in both the converters and the rever-beratory furnace, depending on variations of temperature, matte composition, and reverb gas composition occurring in ordinary plant operation. Within reasonable limits, the field conditions for formation or reduction can be predicted, and probably can be set up and maintained. Converter conditions affecting magnetite formation can be put into numerical values better than for the reverb from purely technical calculations. The converter can be operated so as to keep the magnetite in the slag down to between 12 and 14 pct and still give satisfactory life for the converter brick. This depends upon having converter flux available which will make a slag with a good separation without raising the temperature too high. In the Hurley reverb and others with similar conditions, it is likely that a compromise of conditions will give a reasonably good control of combustion and still keep the magnetite from building up on the bottom. This discussion consists of three main parts. The first is the working out of the equilibrium conditions in the converter for determining in which direction the reaction 3 Fe,,O, (s) + FeS (1) F? 10 FeO (1) + SO, will go under actual converter operating conditions. The second deals with the influence of the solubility of magnetite in the slag and matte in the reverb on the difficulty of tapping matte. The third is an approximation of the equilibrium conditions in the

Jan 1, 1957

By Yang Tsang Woo

The method of making coke that has been adopted at the Kaiping and other collieries in northern China resembles, to some extent, the familiar bee-hive oven process of the United States, except that a kiln with permanent sides is used instead of a closed oven, and the products of combustion are drawn off at the bottom. In building coke-kilns in China, two circular holes, 8 ft. 6 in. in diameter at the bottom and 13 ft. in diameter at the top, are excavated in the ground, and connected at the bottom by a flue, as shown in Figs. 1 and 2. The sides and bottom of the kiln are lined with fire-bricks, and covered with a lager of fireclay, and a rectangular opening, 6 in. by 8 in., connects the bottom of each kiln with the main flue. Around the edge of the hole, and overlapping the lining, is a wall of fire-brick extending 2 ft. 6 in. above the level of the ground. An open space 5 ft. wide is left in this wall, through which the coal is carried into the kiln and the resultant coke removed. This space, of course, is bricked up during the time that the coking-operation is being carried on. The wall is pierced with nine flue-openings uniformly placed around the circumference, each interverliilg space in the wall being strengthened with a brick buttress, as is shown in Figs. 2 and 4. The floor of the kiln, as well as the main flue, is lined with fire-brick. Before charging the coal, it is first placed in shallow bamboo sieves, 2 ft. in diameter, and immersed in a trough 10 ft. long, and 3 ft. wide, fed with pit-water, in which it is washed by hand. Fig. 3 shows the manner in which the washing is done. From 25 tons of good "dust" coal, 22.5 tons of washed coal are obtained. When ready to charge the coal into the kiln, a temporary cone-shaped flue of fire-wood and lump coal is first built to a

Jan 1, 1906

By James P. Kimball

Sulphur is always present in mineral coal of every variety. In the oxidized state it may exist as sulphuric acid in combination with a base. In the unoxidized state it exists in combination with iron as bisulphide of iron. It is also supposed by some chemists to exist in unknown combinations with the organic elements of the coal. This supposition requires the demonstration as a fact that certain coals yield more sulphur than what may exist in a free state, or what belongs to recognized combinations, not only with the accompanying iron but with other inorganic bases abundantly occurring in coals, but which unfortunately are very rarely estimated. This practically as well as theoretically important question it is at present impossible to decide, simply for want of available thorough ultimate coal analyses including the ash. Such analyses, at the best imperfect, are luborious and intricate, and therefore seldom made, the much simpler proximate analyses, so-called, being resorted to by most chemists for technical purposes, and commonly substituted for more scientific work even in the laboratories of government surveys. And it is unfortunate enough that few ultimate analyses are extended to the composition of the ash, much less to its relation to the original constitution of the coal. Sulphur chiefly occurs in most coals in the form of yellow pyrites (pyrite) and white pyrites (marcasite), both minerals, characterized by differences in crystalline form and in physical properties as well as in color, but having the same chemical composition (FeS2). No coal is entirely free from one or both of these dimorphous forms of bisulphide of iron. It may be disseminated through the mass so as to be invisible, in plates or scales between the laminæ of the coal, or in conspicuous segregations or nodules, sometimes of considerable dimensions. Coals greatly differ as to the proportion of sulphur which they contain. They also differ as to the relative proportion in which its several forms of combination occur. Such qualitative and qnantitative disparities arc scarcely, if at all, more marked in separate coal seams of a uniform type than in different parts of one and the same coal seam, whether it be considered locally as constituted of several

Jan 1, 1880

By J. E. Campbell, H. E. Parkinson, Y. P. Chugh

Miner-bolters, which combine mining and bolting functions into a single machine, are being developed by the US Department of Energy to improve production and productivity in coal mining operations. The authors have been studying possible production gains from the use of the Joy automated miner-bolter (AMB) in mine development using the computer simulation model "UGMHS" developed by the Pennsylvania State University. In a recent paper (Chugh, Campbell, and Parkinson, 1979) the authors discussed justification for the development of miner- bolter machines, assumptions and analytical procedures used in computer simulation, cutting, and bolting cycle times for an AMB, and production potential of AMB in four and five entry mine development sections. The effects of available operating time, haulage capacity, bolt installation time, and flush and staggered entry layouts on production potential of AMB were also discussed. In all cases, the production potential of AMB was compared to a full face continuous miner and an independent dual-boom roof bolter operating in a similar mine layout. Present Study In this study, the production potential of the Joy AMB in three and six entry layouts is presented. A three entry mine layout represents development entries in a longwall panel. A six entry layout represents development of main and submain entries in longwall mine development. Flush and staggered entries were considered in both layouts. The assumptions and analytical procedures used in this study were similar to those in the earlier study and are not presented in this paper. Only the results of the studies are presented here.

Jan 1, 1984

By A. Jowett

An attempt is made to explain differences in the size-by-size response to flotation of some sulphide minerals. Physical factors involved in particle-bubble collision and adhesion, and also in possible subsequent disruption of aggregates, are considered especially in relation to flotation of the coarser sizes. Emphasis is on practical observations and their significance, with an essentially qualitative theoretical treatment designed to form hypotheses for further experiments. In particular, the theoretical effects of contact angle and induction time are examined. An argument is developed which suggests a rapid increase in induction time as particle size increases, except for particles small enough to be moving under laminar flow conditions. This can adequately explain poor recoveries of coarse particles in flotation without a need to assume disruption of particle-bubble aggregates in turbulent zones in a flotation cell. Flotation recovery is shown to be very sensitive to small changes in induction time. Consideration of the physics of disruption of particle-bubble aggregates leads to the conclusion that this mechanism does not adequately explain differences in flotation recovery of coarser sizes of some sulphide minerals. For example, to explain differences in recovery of coarse galena and chalcopyrite by this mechanism would necessitate accepting that conditioning with ethyl xanthate which would promote a 60º contact angle at a galena surface would promote only an 11º angle at a chalcopyrite surface. Small differences in induction time arising from differences in surface chemical characteristics of the various minerals offer a more satisfactory explanation of differences in recovery of coarser sizes.

Jan 1, 1980

By John E. Riddell

This review is based on a questionnaire sent out in 1959 to 24 Canadian exploration companies, known to be active in the mineral exploration field, concerning their practice in the use of applied geo-chemical techniques in Precambrian terrains. Of the companies questioned, 21 returned completed questionnaires and of these, 15 companies stated that they had some experience in this type of work. Of the companies who stated that they had no experience in geochemical prospecting in glaciated Precambrian areas it is interesting to note that only three have reported that they do not use applied geochemistry at all. The others have reported that they make use of applied geochemistry in New Brunswick and in the Cordillera in areas of light glaciation. Their opinion is that it is a useful aid when accompanied by other techniques in evaluating mineralized areas. GEOCHEMICAL SURVEYS Table I reviews the type of geochemical surveys used by these companies, and the following points of interest should be noted.

Jan 11, 1960

VARYING in thickness and in number from place to place, coal seams in the Canadian Rockies also range in pitch from nearly horizontal to vertical, sometimes with overturns. Over the entire coal-bearing area there are considerable differences of rank in coals of the same geological age, and there are marked differences in ash content and washability characteristics. Correlation of seams at mining operations within a few miles of each other has often proved impossible. These factors influence mining methods, and, of course, production results. The oldest coal formations of western Canada are of Lower Cretaceous age, as the carboniferous sediments are marine and contain no coal. Coal is also present in formations of Upper Cretaceous and Tertiary ages. The rank of coal varies from semianthracite in some operations in the Kootenay and Luscar formations of the Lower Cretaceous in the Rockies to lignite in the Tertiary fields of the Saskatchewan prairies. Fig. 1 shows the general extent of the formations.

Jan 8, 1954

By H. F. Winham

THE history, development, subsurface geology, production, economics and estimated reserves are discussed in this paper. The Magnolia structure is an anticline with a known maximum structural relief at 372 ft., and the areal extent of the producing acreage is 4494 acres. Production is obtained from the Smackover oölitic lime and 116 successful wells have been drilled. The field, which was developed on 4o-acre spacing, has been under strict control. It is believed to be producing with a strong water drive. To date it has produced more than eighteen million barrels and the estimate of ultimate recovery is 220 million barrels. EARLY HISTORY The Magnolia field is in Columbia County, Arkansas, T. 17 S., R. 19 and 20 W. Various geologists had long been interested in this general area, and as early as 1921, John F. Magale was able to define a surface structure. The first wildcat, one of the first in Columbia County, was drilled in 1923 by Mid-States Oil Co., in sec. 10, T. 17 S., R. 20 W. This well was about 1 ½ miles northwest of the Kerlyn Oil Company&apos;s discovery well and was based on surface geological work done by jewel and Dobie. Subsequently, several shallow wells that were not carried below the Upper Cretaceous beds were drilled in the immediate area of the discovery well, but none of them contained any commercial showings of oil or gas. In the latter part of 1935, Elam and Magale (or Southwood) Garrett No. I was located in sec. 13, T. 17 S., R. 20 W., ¾ mile east of the discovery well. It was drilled to a total depth of 4018 ft. in the lowermost Glen Rose beds but failed to obtain any showings of oil or gas. Interest had been stimulated in this general area, however, by the discovery of several oil fields, mainly Glen Rose oil production in the Rodessa field in Louisiana in 1935, Smackover lime production in Snow Hill in May 1936, production from the Cotton Valley formation in Schuler in 1037, and Smackover lime production from the Buckner field in November 1937. With this background and subsequent seismic work, which was carried on partly because the old shallow wells drilled in the township had shown evidence of closure in the Upper Cretaceous beds, the discovery well, Kerlyn Oil Company&apos;s Barnett No. I, sec. 14, T. 17 S., R. 20 W., was originally projected to a total depth of 6300 ft., to test the equivalents of the Morgan sands of the Cotton Valley series, then producing at Schuler. The well was carried to a total depth of 6325 ft. and tested the objective beds without finding commercial production. In the meantime, however, one well had found production in the Jones sand and another in the Smackover lime at Schuler. Kerlyn immediately made arrangements to deepen the Barnett well to the Smackover lime. The top of the Smackover lime was encountered at 7618 ft. and the well was cored and drilled to a depth of 7740 ft.

Jan 1, 1942

By O. W. Ellis

THE writer has dealt with the effect of various methods of melting copper upon the gas content of the metal.1 The copper referred to in his in his earlier paper was melted in the foundry both in the oil furnace and the electric furnace. As the metal, did not measure up to conductivity FIG. L -INGOT OF OVERPOLEP COPPER. standard, the term "casting copper" was applied to the material in the title of the paper. It was thought that some information in regard to the gases in a sample of overpoled copper might also be of interest, and through the courtesy of C. R. Stevens, superintendent of the copper mill of the Westinghouse Elec. & Mfg. Co., such a sample was obtained. The metal was chill-cast in the form of an ingot about 7 in. long, 2 ½ in. deep (from the base to the raised surface of the ingot) and from 1 3/8 to 1 7/8 in. wide. A photograph of part of the ingot is shown in Fig. 1. The effect of overpoling on the surface of the metal is here quite clearly shown.

Jan 1, 1929

By John E. Willson, Ben L. Seegmiller

A portable electronic instrument was designed and constructed to detect unknown stress magnitudes in rocks. The principle used to detect stress is based on the propagation velocity method. This method allows the stress in a rock subjected to a load of unknown magnitude to be determined, provided that the wave velocity-stress relationship for similar rock is known. This is accomplished by measuring the travel time of longitudinal mechanical waves passing between two points a set distance apart in a rock. The velocity of the wave is calculated and the stress determined from the wave velocity-stress relationship. If the sending and receiving transducer spacing is constant, a time vs. stress relationship rather than the velocity-stress relationship may be used. The method is nondestructive and tests can be made without drilling or otherwise disturbing the rocks. The first studies undertaken in the United States to determine stress in rocks using propagation velocity techniques were reported by Obert1,2 in 1939 and 1940. The Soviet Union first reported using propagation methods to study rock pressures in 1951.8 Success of the method led to the development of a pulse-type ultrasonic seismoscope4 in 1953. Using this instrument, Ivanov and Betaneli5,6 in 1963 succeeded in devising and testing under field conditions a method of investigating coal pillar stresses. In 1967 Osipova7 reported results of similar studies in the Nakhichevan salt mines. In France, Tincelin8 has used the propagation velocity method to study the stress distribution in iron ore pillars. Uhlmann9 has investigated stresses in salt and potash pillars in Germany, using velocity techniques. Determination of uniaxial stress by propagation velocity methods is limited to rocks which have readily detectable wave velocity-stress variations. The present stress detection instrument is restricted in application to rocks which have a velocity change under stress of at least 300 fps. Examples of rocks which meet this requirement are sandstones, coal, and possibly some limestones. Testing of this instrument was limited to a laboratory study and the results may or may not be indicative of what would be found in a field test. A program of field experiments to study the feasibility of using this instrument to determine mine pillar and tunnel stresses is in progress. Instrument Design The instrument has two main components: A probe and a control-display unit. The probe is a hand-held device to which two identical rodlike transducers are rigidly mounted. Coaxial cable connects the probe to the control-display unit which is mounted in an enclosed carrying case measuring 7 % x 9 x 13 in. The instrument is designed so it may be carried and operated by one man. Weight of the probe and enclosed control-display unit is approximately 20.1b. The probe consists of two transducers identical in construction. One transducer is used to convert electronic pulses to mechanical waves and transmit these waves into a rock. The other detects the transmitted waves in the rock and converts them back into electronic signals. The basic element in both transducers is a piezoelectric crystal. The crystal is a disk made of lead titanate zirconate and has a natural resonant frequency of 400 kHz ±1% in the thickness mode. A schematic of one of the two identical transducing elements is shown in [Fig. 1.] A spacing of 6 in. between sending and receiving transducers has been found to be most satisfactory. The control-display unit consists of a pulse generator and a 1-in. oscilloscope. Various electronic devices are used for the power supply, amplification, and calibration. The amplitude of the square wave from the pulse generator can be continuously varied between 0 and +20 vdc. Pulse width can be set at 1, 2, 3, or 4 µsec. The repetition rate of the pulse can have the following values: 50, 75, 100, 150, 200, 250, 300, 400, 500, 750, and 1000 Hz. The oscilloscope delay system allows the travel time of the longitudinal wave passing through a rock to be measured to an accuracy ±0.1 µsec in the range from 0 to 995 µsec. Fig. 2 is a schematic of the various electronic sections in the control-display unit. Laboratory Testing The first step in using the instrument is to develop a velocity-stress or a time-stress curve for the par-

Jan 1, 1972

By W. F. Mullen, C. W. Stickler

WITH few exceptions, unit operations in the Pennsylvania slate industry in 1950 did not differ appreciably from production methods described by Behrel and Bowles2-4 several decades ago. Many traditionally picturesque but relatively inefficient hand operations continued to contribute to high operating costs in an industry in which the margin of profit, for various reasons, was admittedly low. As part of a general program in which the Pennsylvania State College has been assisting the slate industry to solve some of its problems, time and method studies were used in an effort to determine the bottlenecks of production flow. The results are as applicable to other stone industries as they are to slate. A study of the basic elements in the production of roofing slate, structural slate, and slate blackboards among selected slate producers revealed comparatively few refinements. Any marked difference in operating method was characterized largely by the available equipment, a noticeable difference in the working properties of the rock quarried by a particular company, and by either the habit thinking or ingenuity of the individual operators. With the hope that significant economies in manufacture might result, all the unit operations were broken down into elements for study and analysis. The greater part of the study encompassed those operations from the quarry rim to the finished product; however, to obtain a reasonable synthesis of time allotments per operation, several analyses of quarrying from the finish of the wire saw cut are included for reference. The complete study of quarrying made by Bowles several years ago resulted in the introduction of the wire saw into the operation; there appears to be a need for as fundamental a change in the processing of finished slate as the wire saw was in the quarrying process. Fig. 1 illustrates the flowsheet of quarried rock from the parent bed to the quarry rim with the basic elements which contribute to the difficulty of removing slate from the ground. In Pennsylvania roofing slate is produced by one of three procedures, many operations of which are quite similar. In the traditional or classical method the quarried rock is reduced by sculping and auxiliary manual treatment to workable size and then is split and dressed (trimmed to size) in the conventional shanty. Archaic as this method might appear, there are several elements that result in substantial savings over more mechanized operations. In a modified version of this method, the block is reduced by sawing and gouging preparatory to splitting and dressing in indoor stations. This method is particularly adaptable to rock which fractures unevenly by conventional sculping methods. In the third and more modern method, reduced block received from the block maker is diamond-sawed to length and finished on a production line. Fig. 2 illustrates the flow pattern in each case. In the modified and more modern plans, dressing is done on mechanically-operated trimming machines, which appreciably reduce the fatigue factor of the operator. However, it should be noted that it is advantageous for the operator to be able to control the speed of the knife to prevent breakage of certain classes of stock. In the classical plan the splitter and dresser normally act as blockmakers also and carry their eighter slabs (equivalent in thickness to eight shingles) into the shanties themselves, all of which adds to fatigue and reduced efficiency. The extremely low investment and operating expense of the classical method has undoubtedly been of paramount importance in its continuation. Production of Mill Stock The production of mill stock is accomplished by less picturesque methods than enter into the production of roofing slate but it is, in the main, relatively more efficient because of the increased use of machinery for handling and for finishing. Mill stock can be classified into two categories as far as difference in production method is concerned: structural slate and blackboards. The significant difference in operation is caused largely by the nature of the rock used for each product with the best and easiest cleaved rock being reserved for blackboards to permit the splitting of large slabs to as little as 1/2 in. in thickness. Structural stock, on the other hand, is split into panels no thinner than 11/2 in. and

Jan 1, 1952

By George A. Packard

The shaft of the Raven mine, at Butte, Mont., is an incline 1,700 ft. in length and dipping at various angles. At the top the dip is 70" from the horizontal, but this is gradually flattened until at the 300-ft. level the inclination is only 47. This angle continues to the 1,100-ft. level, below which it curves with a 125-ft. radius to 78". In addition to these variations in dip, the shaft does not lie in the same vertical plane, with the result that the hoisting rope not only rubs at intervals on both the hanging and foot walls, but presses strongly against the west dividers near the collar, while 300 ft. below it runs close to the east end plates. The early operators used no idlers, aid wall plates cut 6 in. deep by the rope resulted. Later operators first attempted to overcome the excessive friction, and the wear of rope and wall plates, by introducing solid cast-iron idlers, 3 in. in diameter. To allow for the travel of the rope from side to side some of these had to be 3 ft. long and were very heavy. Judging from the appearance of the old idlers of this type found at the mine, they often failed to turn in the bearings, which is not surprising when it is considered that they would have to make 1,000 rev. a minute under ordinary hoisting conditions. The next rolls were made of wood, 6 in. in diameter, with an iron band about each end, and a pintle of 1-in. round steel driven in at the ends to serve as a shaft. These wore rapidly, and were soon replaced by rolls made from water pipe, 5 or 6 in. in diameter, cut to the desired length and fitted with a wooden cylinder into which the pintle was driven. Where the idlers were used on the hanging wall of the shaft the original bearing was simply a piece of + by 1 3/2 in. strap iron, 10 in. long, turned up at the end in a circle 11/8 in. in diameter to receive the shaft. A small hole served for oiling, and common black oil was generally used, although filtered oil from engine bearings and compressor bearings was also used. When the rolls were to be placed on the foot wall the bearings were made from two pieces of 1 by 3 in. steel, 6 in. long. A half circle was cut on the flat side of each piece and the two half circles together formed a bearing. Oil holes were provided, and in some cases holes were bored through the two pieces so that they could be screwed or spiked to the wall plates.

Jan 1, 1915