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By F. A. Linforth, W. B. Daly

The sampling and estimating of the orebodies in the mines of the Anaconda Copper Mining Co., at Butte, have been made a part of the work of the geological department. The wisdom of this assignment is apparent, considering the organization of that department, the nature of its records, and the scope of its regular work. In the first place, the district is divided into units of two or three mines each and a geologist is assigned to each group. His trips to all the working places are frequent and he follows the mining operations closely. His notes are graphic and detailed. In the office, he posts his notes on the standard maps and sections and on special maps for correlation purposes. With the approval of the chief or assistant chief geologist, he recommends development work underground and watches its progress. This gives the geological department a precise knowledge of the character of each orebody. Sampling is done by one or two samplers at each mine working under a chief sampler stationed in the geological office. The chief sampler has access to all the office maps and is in frequent consultation with the geologists; his instructions to the samplers are issued with the approval of the chief or assistant chief geologist. Not only is it important that the geologic information be used to direct the sampling but the reports of the samplers and the records kept by their chief are useful to the geologist, who needs them in his work of estimating ore reserves and deciding on recommendations for development work. The sampler, moreover, is at the mine every day and makes a daily report, which keeps the geologist in close touch with the progress of development work. The sampler is attached to the mine office and thus comes under the jurisdiction of the operating department. In fact, any special orders prepared by the chief sampler are submitted to the general superintendent, after approval by the geological department. The mine foreman may

Jan 1, 1923

By Harold H. Haman

In view of Twentieth Century advances in mineral dressing technology, most modern metallurgists would be inclined to class jigging as an early, somewhat primitive, representation of their art. The practice of metallic sulfide concentration by jigging is becoming rare. This is so, in part, because the type of ore amenable to the process is becoming rare. It is appropriate, however, that this symposium recognize that a significant contribution of domestic zinc remains dependent upon coarse preconcentration by jigging, and that the state of the art be described even though the theory and applications are already well delineated in technical literature. One of the few areas still using jigging as a major concentrating process is The Upper Mississippi Valley. AREA HISTORY AND JIGGING The Upper Mississippi Valley (UMV) has produced one billion dollars worth of zinc and lead at current prices. The district includes the southwest part of Wisconsin, the northwest corner of Illinois and an adjoining narrow strip of Iowa. Within a total area of 4,000 square miles, important sub-districts cover about 400 square miles. As is widely known, early mining in this district involved shallow lead---crevice deposits of galena that were very selectively mined, hand sorted and cleaned, and then smelted at or near the site. The district has produced almost a million short tons of metallic lead, a very large part of this during the fifty- year period following 1830. Zinc mining began at the end of the civil war, but at a very low rate until about 1905.

Jan 1, 1970

By James Douglas

No one mill attribute it to the partiality of friendship if I claim for Thomas Sterry Hunt the right to a place among that little band of men whom nature has endowed with what seems an almost magical power of insight into her mysteries. But those of us who knew him well, while me never ceased to wonder at the breadth of his genius, the fertility of his imagination, and the wonderful capacity of his memory, learned that these native powers would have availed little to win him his rank in science, had they not been stimulated by ceaseless mental activity, indomitable perseverance, and unflagging industry. Thomas Sterry Hunt was born in Norwich, Connecticut, on September 5, 1826. He came on his mother's side from that good old stock which gave to Puritan England the gentle mystic Peter Sterry and that uncompromising preacher Thomas Sterry, who wrote the notable tract, " The Rot among the Bishops " in 1667, and that gave to New England Consider and John Sterry, the mathematicians. While lie was a child the family moved to Poughkeepsie on the Hudson. There the father died when Thomas, the oldest son, was twelve. The mother and her family of six young children returned to the old home in Connecticut. For a short time Thomas attended the public school; but it was for a short time only, as he, thus early, was required to share the burden of the family support and to seek employment. This he found first in a printing-office, then with an apothecary, afterwards in a book-store. The surroundings may have been congenial, but the occupation probably was not; for he re-

Jan 1, 1893

INTRODUCTION TO MINERAL INVENTORY The first explanations to be made must answer the questions: What is a mineral inventory, and how is a mineral inventory different from an ore reserve? The term mineral inventory is a relatively new designation-born of the computer age and grown to common usage with the development of computer techniques for extending drill-hole sample values to regularly gridded data blocks representing in situ mineral values. The term has been generally accepted by geologists and engineers because it has a precise meaning which avoids many of the ambiguities of ore reserve. A mineral inventory is an estimated inventory of mineral in place and is usually achieved by dividing the deposit into regularly spaced gridded blocks to which are assigned estimated values using various extension techniques. A mineral inventory makes no presumptions about the mineability of the blocks or their cutoff grade. Such parameters are implied with the use of the term ore reserve. Ore has a specific definition (see Chapter 1) and ore reserve must reflect that definition and thus has economic and mining constraints. An ore reserve will be different under different economic and mining conditions and is subject to reclassification by many factors. The term mineral inventory avoids all such ambiguities. A mineral inventory estimate may be good or poor according to how much is known about the deposit, how extensively it has been sampled, and what techniques have been employed to assign estimated grade to the blocks. Such an inventory, however, does not change with time, economic conditions, or mining technology, but only with new and better data and methods. Thus, the mineral inventory is an excellent designation for describing mineral value distribution within a deposit's geological setting. The techniques for assigning values and tonnage to mineral blocks or panels are the subjects of this chapter. Whether a summation of the blocks be called a mineral inventory or mineral reserve is not important. Many conventional methods for computing reserves of minerals refer to the methods in terms of ore reserve calculations. OBJECTIVES OF MINERAL INVENTORY OR RESERVES Mineral inventories are computed for the purpose of determining the quantity, the quality or value, and the spatial distribution of potentially economic minerals within the earth. Such computations are made at all stages in the life of a mineral extraction process, from the early exploration phase to the final year in the life of a mine. To quote Popoff (1966, p. 2), "They are the most responsible and irreplaceable tasks in the valuation of a mineral deposit. Efficiency in extraction and productiveness is impossible without accurate reserve computations." The accuracy of estimate required for each of the three parameters of quantity, value, and location is dependent upon the objective for which the reserve estimate is being made. Again quoting Popoff (1966, p. 2): Reserves are computed to determine the extent of exploration and development; distribution of values; annual output; probable and possible productive life of the mine; method of extraction and plan design; improvements in extraction, treatment, and processing; and requirements for capital, equipment, labor, power, and materials. Such computations are used to assist development

Jan 1, 1980

By T. S. Fuller

VARIOUS phases of the embrittlement of solid copper containing oxygen by the action of reducing gases at high temperatures through the work of many experimenters are familiar to readers of metallurgical literature. In 1900, Heyn1 recorded observations of copper being heated in hydrogen with the development of fine cracks, the result of the formation of steam within the metal through the action of hydrogen on the copper oxide usually found in copper. Archbutt,2 in 1905, published the results of experiments, showing that all oxygen can be removed from copper by heating to redness, for several hours, in hydrogen. Intercrystalline cracks and extreme brittleness developed. Archbutt attributed the brittleness to the reduction of oxides present in the copper. Bengough and Hill,3 in 1910, encountered the now well-known embrittlement at the outset of their investigation and stated in their paper: It soon became evident that the effect of any heat treatment depended largely on the presence or absence of reducing gases in the atmosphere of the furnace, as well as on its temperature. Hence it became necessary to adopt a somewhat elaborate scheme of heat treatment. The authors found embrittlement in separate experiments due to both hydrogen and carbon monoxide and concluded that: "The gases ruin the alloys after being heated to about 650° C. in their presence."

Jan 2, 1926

C. A. HANSEN (communicated appendix*).-Since the above paper was written, tests have been conducted with a view to securing a sounder basis for discussing the effects of temperature on the behavior of the zinc cell. The following data, while making no pretense to great accuracy, are still reasonably close. Cells were operated with the same feed solutions, but at varying temperatures, the latter being controlled by means of immersed steam coils. Current efficiencies obtained indicate that the corrosion rate

Jan 11, 1918

By D. N. Stevens, D. N. Bloom, M. A. Klugman

This paper represents a portion of a larger study concerning dispersion of major, minor and trace elements in wall-rock about orebodies in part of the Colorado Mineral Belt. This phase of the work pertains to the Mg/Ca ratio variations in wall-rock adjacent to orebodies in two mines, one in the Horseshoe Mining District (Hilltop mine), and the other in the Alma Mining District (Phillips mine). The two districts lie adjacent to one another, the Horseshoe District on the south and the Alma District on the north, and are commonly referred to as one, the Alma-Horseshoe Mining District. The Alma-Horseshoe Mining District is located 80 miles southwest of Denver, Colo., (Fig. 1) in Park County, on the eastern slope of the Mosquito range. The area is typified by rugged relief with altitudes of from 10,500 to 14,000 ft. The Horseshoe District is separated from the Lead-ville Mining District by the crest of the Mosquito range, with the Leadville District to the west. The Alma-Horseshoe District has been considered as peripherial to the Leadville District, and has not received the detailed attention that the Leadville District has. These two mines were selected for this study because of their ease of access; the attitude of the workings and the drill holes to the known mineralization; the similarity of the mineralization; and because the mineralization occurs in carbonates in both mines but are of different ages. In the Hilltop mine mineralization is confined to the Leadville dolomite (Mis-issippian) and is principally lead with subordinate zinc and silver. In the Phillips mine the mineralization occurs in the Peerless dolomite and shale (Cambrian) and in the Manitou dolomite (Ordovician) and is principally zinc with subordinate lead, silver,

Jan 1, 1964

By Ernest R. Dickie

BRIEFLY, the ore body of the Bagdad Copper Corp., Bagdad, Ariz., is a monzonite porphyry carrying copper values fairly evenly distributed from the surface down through the primary zone. Tabular in shape, the ore body extends over several hundred acres. The oxidized zone averages about 60 ft thick and assays about 0.50 per cent copper. Under this is the chalcocite zone, or the zone of secondary enrichment, which averages about 100 ft in thickness and assays about 1.4 per cent copper. Under this zone lies the primary zone which extends in places to a known depth of about 150 ft and averages in grade about 0.60 per cent copper. Several years ago the block-caving method of extracting the ore from the chalcocite zone was installed. Then it was necessary to draw only about 250 to 300 tons per day, the capacity of the old milling plant. Early in 1943 the new 2500-ton concentrator was put into production. Drawing ore from the developed stopes in sufficient tonnage to furnish the mill at capacity proved unsuccessful in more ways than one. First, drawing the ore at an accelerated rate caused a large amount of dilution, there-by lowering the grade, and also caused funneling through to the surface. The surface material was detrimental to milling metallurgy and development of these stopes was costly per ton of ore extracted. The grade of ore decreased. making it impossible to maintain a mill head of over 0.9 per cent copper. The

Jan 1, 1947

By John M. Kelly

This report covers the development, exploratory drilling, and production of oil and gas in New Mexico during 1945. The statistical information was gathered from state regulatory agencies and other sources believed to be reliable; it is presented here in combined form for general information only, and the author assumes no responsibility for its accuracy or completeness.

Jan 1, 1946

By John Wulff

IN a previous paper Burwell and Wulff1 have shown by electron diffraction studies that allotropic transformations can be induced in 18-8 stainless steel by polishing to a depth of about 5 X 10-5 cm. The development of a successful etching technique, which removes accountable amounts of metal from the surface without pitting and without changing the root-mean-square profile as measured with an Abbot profilometer, led to the application of the whole technique and material to an investigation of various mechanical methods of surface preparation, such as grinding, lapping, superfinishing and cold-rolling. Ideally suited to such a study is 18-8 of low carbon and nitrogen content. The material can be heat-treated to produce a coarse-grained single-phase structure. On cold deformation there occurs not only a reduction in grain size but, at regions of maximum stress, a change of face-centered material to body-centered material, provided the metal is worked below about 200°C. When worked between 200° and 415°C., diffraction studies evidence a reduction in grain size without the appearance of any new body-centered ferrite. Ferrite, which can be referred also to strains induced by carbide precipitation, is evident when the material is worked at temperatures between 415° and 800°C. Stainless steel is advantageous because the usual surface-finishing operations do not readily oxidize it to any appreciable depth. In order to produce detectable oxide films on the surface of austenitic 18-8 in a time equivalent to that required to produce, for example, a fine ground finish on ½ sq. in., the sample must be held in a furnace or oxygen stream above 600°C. By using longer heating times, it is possible to achieve measurable oxide films on samples whose surface temperatures exceed only about 250°C. EXPERIMENTAL PROCEDURE The apparatus, material, and technique described in the previous paper1 were employed in this investigation. Samples of 18-8, 1 by 1 cm. by 4 mm., were carefully annealed in the austenitic range to produce a coarse-grained structure. They were then quenched and pickled to remove scale. On a lapping machine of the Graton2 type, both sides of each sample were polished to a flatness of 1/6 a wave length of sodium D light. Thereafter they were electrolytically etched to remove at least 1 X 10-3 cm. of the surface, a depth that is well below that of any deformations induced by surface preparation. Samples of this material, carefully mounted in steel holders, were then sent to various industrial organizations known for special finishing techniques. In the instrument shops at the Massachusetts Institute of Technology check samples were finished by the same methods, as far as they were available. Abbot profilometer3 readings were made on all the samples, which were then carefully degreased in double distilled benzine and alcohol, dried in a hot-filtered air blast, and kept in a desiccator. Suc-

Jan 1, 1941

By Howard N. Eavenson

Waste of coal, or perhaps more properly the percentage of its recovery in mining, has keenly interested me during an experience of over a half century in coal mining. In the early part of that time and until a few years after World War I the use of better mining plans and supervision resulted in an increased recovery. Since then, and particularly since controlled prices have been effective and mechanized mining has been extensively practiced, the tendency has all been toward lesser recovery in mining. The purpose of this paper is not to criticize the methods being used, which after all are the result of economic forces, nor to point to any particular mining method or to any locality where the best recovery is obtained, but to call attention to the tremendous waste of our national coal resources that is constantly continuing. A possible method of reducing this loss will be suggested, but without any illusions that it will be followed or that such attention will be paid to it. Average Recovery The actual average percentage of recovery of coal in the ground is not known. Many papers giving figures of recovery have been published but no definite base of measurement has been agreed upon by those presenting the data. As an example, the Pittsburgh bed can be considered—it has produced a greater output than any other and has been sold or leased at higher prices in place than any other coal bed. Over a large part of its area where high prices have prevailed it has been customary for many years to draw pillars. In these mines it has been usually assumed that the average recovery was higher than in other beds and percentages of recovery of 85 to 95 per cent are usually said to be realized. In all of these mines, however, it has been customary to leave from 4 to 8 in. in the bottom, and 10 to 16 in. in the roof; thus from 14 to 24 in. of coal has been left in the mines out of a total seam thickness of about 96 in., or 15 to 25 per cent of the total coal. No attempt is made to mine this coal for metallurgical purposes, owing to its impurities, although this unmined coal is of excellent quality and better than much of the coal used for steam purposes. If we add to this discard the coal usually lost in mining and those parts of the bed 'that must be left because of bad seam conditions, barrier pillars, reservations for buildings, gas and oil wells, one can easily see that the actual recovery from this bed cannot be more than 65 to 75 per cent. In addition, the rooster or rider or wild coal over the mining bed—often 3 to 4 ft. of coal, high in ash and sulphur but still good steam coal—is hardly ever mined, nor is any attempt made to mine it. It is safe to say that from a broad economic viewpoint not more than 50 to 60 per cent of the fuel value of this great seam is actually being recovered.

Jan 1, 1947

By Howard N. Eavenson

Waste of coal, or perhaps more properly the percentage of its recovery in mining, has keenly interested me during an experience of over a half century in coal mining. In the early part of that time and until a few years after World War I the use of better mining plans and supervision resulted in an increased recovery. Since then, and particularly since controlled prices have been effective and mechanized mining has been extensively practiced, the tendency has all been toward lesser recovery in mining. The purpose of this paper is not to criticize the methods being used, which after all are the result of economic forces, nor to point to any particular mining method or to any locality where the best recovery is obtained, but to call attention to the tremendous waste of our national coal resources that is constantly continuing. A possible method of reducing this loss will be suggested, but without any illusions that it will be followed or that such attention will be paid to it. Average Recovery The actual average percentage of recovery of coal in the ground is not known. Many papers giving figures of recovery have been published but no definite base of measurement has been agreed upon by those presenting the data. As an example, the Pittsburgh bed can be considered—it has produced a greater output than any other and has been sold or leased at higher prices in place than any other coal bed. Over a large part of its area where high prices have prevailed it has been customary for many years to draw pillars. In these mines it has been usually assumed that the average recovery was higher than in other beds and percentages of recovery of 85 to 95 per cent are usually said to be realized. In all of these mines, however, it has been customary to leave from 4 to 8 in. in the bottom, and 10 to 16 in. in the roof; thus from 14 to 24 in. of coal has been left in the mines out of a total seam thickness of about 96 in., or 15 to 25 per cent of the total coal. No attempt is made to mine this coal for metallurgical purposes, owing to its impurities, although this unmined coal is of excellent quality and better than much of the coal used for steam purposes. If we add to this discard the coal usually lost in mining and those parts of the bed 'that must be left because of bad seam conditions, barrier pillars, reservations for buildings, gas and oil wells, one can easily see that the actual recovery from this bed cannot be more than 65 to 75 per cent. In addition, the rooster or rider or wild coal over the mining bed—often 3 to 4 ft. of coal, high in ash and sulphur but still good steam coal—is hardly ever mined, nor is any attempt made to mine it. It is safe to say that from a broad economic viewpoint not more than 50 to 60 per cent of the fuel value of this great seam is actually being recovered.

Jan 1, 1947

By E. H. Rose

The question is often pondered whether the flotation process is still an art or has become a science. The fact is that flotation is a science in so many variables that only art can blend them. It is demonstrable that no less than one hundred variables are involved in any flotation operation, most of them immeasurable on a routine basis. The behavior of every one of them is always governed by orderly and unchanging law, as no man can doubt. There is nothing truly indefinite about that behavior; the only indefiniteness is in our knowledge of it, either in general or from hour to hour in the circuit. In dealing with a simultaneous equation in, say, one hundred variables, all not only affecting the outcome to some degree but many influencing each other as well, it is small wonder that the mineral-dressing profession presents a case history of debate and controversial opinion. The statement of the subject of this symposium—a symposium on flotation machines—seems to delimit and define the matters to be considered and at least the general nature of the conclusions sought. Nevertheless, in such a meeting it is not possible to meditate upon all the variables that do in one way or another affect performance of flotation machines. To attempt it would be much like trying to operate a hundred-ring circus. Perverse variables being what they are, one could expect the elephant in ring 36 to squirt a trunkful of water at the tightrope performer in ring 37 to ruin that artist's act, just as our learning to depress pyrite in the early 1920's ruined the act of the formerly faithful old Callows. Honest Disagreements Selective flotation dates from 1922 or 1923, when several plants established their regular flotation operations on a selective basis. Since then, two decades of efforr have gone into the development of better and better flotation machines. Many have come and gone. The competition has simmered down to perhaps a dozen machines, each one of which is staunchly defended by some operators as being the best for their particular jobs, and as scornfully dismissed by at least a few others elsewhere as being positively the poorest for theirs. These disagreements are honest disagreements, honestly arrived at after plenty of sweat. In many a flotation plant, one machine has been tried out against another under conditions as fair as they could possibly be made. Sometimes the results are on a par but more often one machine shows a consistent and unmistakable margin of superiority over the other. In equally honest and enlightened tests, machine A may win over machine B in one place, and machine B over machine A by as great a margin in another. This paper purposes to identify some of the more prominent reasons why this is so,

Jan 1, 1947

By E. H. Rose

The question is often pondered whether the flotation process is still an art or has become a science. The fact is that flotation is a science in so many variables that only art can blend them. It is demonstrable that no less than one hundred variables are involved in any flotation operation, most of them immeasurable on a routine basis. The behavior of every one of them is always governed by orderly and unchanging law, as no man can doubt. There is nothing truly indefinite about that behavior; the only indefiniteness is in our knowledge of it, either in general or from hour to hour in the circuit. In dealing with a simultaneous equation in, say, one hundred variables, all not only affecting the outcome to some degree but many influencing each other as well, it is small wonder that the mineral-dressing profession presents a case history of debate and controversial opinion. The statement of the subject of this symposium—a symposium on flotation machines—seems to delimit and define the matters to be considered and at least the general nature of the conclusions sought. Nevertheless, in such a meeting it is not possible to meditate upon all the variables that do in one way or another affect performance of flotation machines. To attempt it would be much like trying to operate a hundred-ring circus. Perverse variables being what they are, one could expect the elephant in ring 36 to squirt a trunkful of water at the tightrope performer in ring 37 to ruin that artist's act, just as our learning to depress pyrite in the early 1920's ruined the act of the formerly faithful old Callows. Honest Disagreements Selective flotation dates from 1922 or 1923, when several plants established their regular flotation operations on a selective basis. Since then, two decades of efforr have gone into the development of better and better flotation machines. Many have come and gone. The competition has simmered down to perhaps a dozen machines, each one of which is staunchly defended by some operators as being the best for their particular jobs, and as scornfully dismissed by at least a few others elsewhere as being positively the poorest for theirs. These disagreements are honest disagreements, honestly arrived at after plenty of sweat. In many a flotation plant, one machine has been tried out against another under conditions as fair as they could possibly be made. Sometimes the results are on a par but more often one machine shows a consistent and unmistakable margin of superiority over the other. In equally honest and enlightened tests, machine A may win over machine B in one place, and machine B over machine A by as great a margin in another. This paper purposes to identify some of the more prominent reasons why this is so,

Jan 1, 1947

By Paul H. Price, John P. Nolting

The history of the salt industry in West Virginia dates back nearly two hundred years; howTever, the history of salt as an important raw material for the chemical industry is much more recent. The earliest record of West Virginia salt was in 1753, when a raiding party of Shawnee Indians attacked frontier settlements in Virginia taking captive a number of the pioneers. One of these, who later escaped and returned to her friends, described how the Indians camped at a salt spring on the banks of the Kanawha River and evaporated the water to obtain a supply of salt to take back to Ohio with them. Settlement of the Kanawha Valley began in 1774, that same year the Battle of Point Pleasant ended the Indian's power in the Ohio Valley; and settlement of the Kanawha Valley progressed rapidly with a consequent increase in the importance of the Kanawha Licks as a source of salt. In those days the evaporation of the brine was a family matter for which the family wash-kettle was commandeered. In 1797, Elisha Brooks erected the first salt furnace in the Kanawha Valley. It consisted of two dozen small kettles set in a double row, with a flue beneath, a chimney at one end, and a fire bed at the other. He obtained his brine at a depth of 20 to 30 ft by sinking hollow logs, called "gums" into the salt lick and dipping the brine with a bucket. His production was about 150 lb a day and sold for 8 to 10 cents a pound. In 1806, the Ruffner brothers, David and Joseph, prospected for and drilled the first well in America using a spring pole and steel chisel bit. After several failures they succeeded, in 1808, in drilling to a total depth of 59 ft where they secured a good flow of strong brine. Most standard drilling tools, jars, casing, and practically all the basic oil well drilling machinery now in use, were developed in the drilling of salt wells. To evaporate this brine, the brothers built a wood-burning furnace similar to that of Brooks but with a better arrangement and with more and larger kettles. When production was started they reduced the price of salt to 4 cents a pound. By 1817, there were some 30 salt furnaces in the Kanawha area. It was in this year that bituminous coal first was used as an industrial fuel. David Ruffner again pioneered the way by converting his furnace for the consumption of coal. This year also brought about the formation of the first trust which controlled production and set a price of 1 cent a pound for the salt.

Jan 1, 1950

By Charles Williams

BOYLE'S law states that the volume of a gas at B constant temperature varies inversely as its pressure, PV = C, or V = C/P Therefore, the volume of a gas at any pressure is to its corresponding pressure as the volume of the same gas at atmospheric pressure is to atmospheric pressure, V, = V° Also, P1 P0 the volume of the gas at any other pressure is to its corresponding pressure as the volume of the gas at atmospheric pressure is to atmospheric. pressure, V2/P1 V0/P0 V°P VA Solving for V1 and V2, V2 = and V2 =V0P2/P0

Jan 2, 1927

Mining Methods GERALD F. SHERMAN, Chairman JOHN A. CHURCH, Vice-chaiman CHARLES W. WRIGHT, Secretary Alluvial Mining CLINTON BERNARD CHARLES JANIN H.R. NORSWORTHY Open-cut Mining Iron MAX H. BARBER J. A. MACKILLICAN Copper MICHAEL CURLEY R.B. TEMPEST Underground Mining Iron C. H. BAXTER TENNEY C. DESOLLAR

Jan 1, 1934

By W. D. Robertson, R. E. Reed-Hill

Slip on (11221 planes, which do not contain a close-packed direction, has been identified on crystals strained at —190°C. Evidence is also presented to show that (1071) pyramidal slip is a mode of deformation at 25ºC in a limited orientation range. PYRAMIDAL slip (l011) <1120> in magnesium at elevated temperatures has been established for some time.1-2 More recently, Burke and Hibbard reported this mode of deformation at 25ºC, but their results were questioned by Craig.&apos; Also, as far as the present authors know, slip in magnesium on pyramidal planes of the second order, Type II (1122)) has not previously been reported, although Gilman"," identified (1122) slip in another hexagonal metal

Jan 1, 1959

By R. H. Richman, II Conard G. P.

AS part of a study of deformation twinning in bcc crystals, solid solutions of beryllium in iron have been found to twin profusely when strained slowly at room temperature. This note reports some crystal-lographic aspects of the operative mechanical twin modes in a quenched Fe-5 wt pct Be alloy. After melting twice under a partial pressure of helium, the alloy was homogenized, cut into pieces 7 to 10 mm on an edge, solution treated at 1130°C,1 and quenched. Debye-Scherrer patterns confirmed

Jan 1, 1963

By W. A. Backofen, M. L. Ebner, D. H. Avery

Slip on (11221 planes, which do not contain a close-packed direction, has been identified on crystals strained at —190°C. Evidence is also presented to show that (1071) pyramidal slip is a mode of deformation at 25ºC in a limited orientation range. PYRAMIDAL slip (l011) <1120> in magnesium at elevated temperatures has been established for some time.1-2 More recently, Burke and Hibbard reported this mode of deformation at 25"C, but their results were questioned by Craig.&apos; Also, as far as the present authors know, slip in magnesium on pyramidal planes of the second order, Type II (1122)) has not previously been reported, although Gilman"," identified (1122) slip in another hexagonal metal

Jan 1, 1959