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By Nicholas J. Grant, Rolf Nordheim, Thomas B. King

Constant-load creep-rupture tests were conducted in an argon atmosphere at temperatures of 1600 to 2200°F on two heats of iron containing about 0.06 and 0.09 pct P in solid solution. The tests were conducted in the as-cast condition. Stresses were selected to give strain rates approaching those encountered in the hot working of steel, and within the limits of the creep-rupture test apparatus, namely. from about 0.001 to 50 pct per sec. FOR generations, rolling and forging practices have been developed largely by empirical methods, based on prior experience. While largely successful, difficulties are frequently encountered when new or different alloy compositions are tried for the first time. There is need to know the relationships which exist among temperature, strain rate, composition, and other variables on the deformation and fracture characteristics of steels. It has been shown for aluminum and several high-temperature alloys that the deformation and fracture characteristics follow rather well-defined laws and patterns which are closely related to the effects of strain rate, temperature, and composition.1"3 As the melting temperature of aluminum or its alloys was approached, it was possible to study the interplay of work-hardening factors and recovery or recrystal-lization, the latter resulting in improved ductility and increased resistance to intercrystalline fracture. In particular, the effects of composition are very poorly known. Within the meaning of "composition" should be included impurity elements, such as sulfur, phosphorous, lead, copper, and many others, as well as those elements which are added for alloying purposes in iron. The role of multiphase structures, in contrast to solid solutions, should also be studied in much greater detail. high strain rate creep-rupture tests would provide information regarding the deformation and fracture characteristics of metals at hot-working temperatures. It was also desired to establish the stresses which are necessary to accomplish the desired deformation, and to relate these to the strain rate and temperature. Secondly, it was the purpose of this particular study to determine the effects of phosphorus, within solid solution limits, on the deformation and fracture characteristics of iron at hot-working temperatures. EXPERIMENTAL PROCEDURE Two phosphorus levels were selected, one corresponding to the upper limit normally found for commercial steels, and one substantially higher than this, but within the region of solid solubility at the test temperatures chosen for this investigation. Twenty-pound heats were prepared by induction melting electrolytic iron in air in a magnesia lined furnace. After deoxidation with about 0.2 pct Al, ferro-phosphorus was added, and the heats were then poured into prefired silica molds. The purpose of using silica molds was to obtain a slow cooling rate during solidification to produce a relatively coarse grain size, producing a structure subject to some segregation and therefore more nearly characteristic of ingot structures. The major part of the casting assembly provided a large shrinkage head in order to produce sound castings and to promote slow cooling of the specimens. The primary austenitic grain size varied from ASTM No. 2 to No. 00. The composition of the heats, including a low phorphorus melt, used for comparison purposes, is shown in Table I. It will be noted that the oxygen content of the phosphorus containing heats is relatively high, in spite of the deoxidation practice, a condition which is difficult to control in small melts. Creep-rupture tests were nevertheless undertaken when it was noted that this quantity of oxygen did not appear to give results any different from those from a heat containing 0.01 pct 0, low phosphorus and comparable carbon Content;4 see Table 11.

Jan 1, 1961

By G. S. Coe, G. F. Bolling

The columnar-to-equiaxed transition (CET) has often been studied as a function of concentration, C, temperature gradient in the melt, G, and rate of solidification, R. Although it may not be immediately obvious it is sure that the CET, expressed as some ratio G/R necessary to suppress equiaxed nucleation for a given C,' cannot be a single-valued function of C. This restriction which is ignored in theoretical presentations is most trivially illustrated by the following simple reasoning. An ordinary casting is an unconstrained system where the thermal conditions change continuously with progressive solidification, and thus where the ratio G/R decreases from the outset. A CET occurring early in a casting (much equiaxed material) must then mean a higher value of G/R at transition, in comparison with the late CET which implies a more columnar ingot. Therefore, when we choose to experiment under constant superheat for all alloys across the simplest binary, and we find an inversion in the amount of equiaxed material, we shall find two values of C satisfying the G/R condition. However, in addition to this simple instance, the early experiments of ~orthcott' seem to show that the amount of equiaxed material in cast ingots may be a multivalued function of composition in a given binary system. In the present investigation, the CET has been examined in Pb-Sb, Pb-Ag, Pb-Sn, and Sn-Pb under several types of casting conditions. First, we have used the "bidirectional" ingot section Simulation;~ and we have used .'i~z situ" ingot solidification where a hot mold already filled with liquid is chilled so that there is no disturbance provoked by pouring. Differences in method are trivial to our purpose except for the fact that certain variations are more simply observed by one technique than by the other. Second, we have varied the mode of superheating from the simple extreme of casting all alloys from the same temperature regardless of composition to the careful casting at a standard superheat, calculated with respect to the alloy liquidus. Third, we have made provisions for decanting and interface observation, so that some "castings" were not completed. The results obtained with Pb-Sb are exemplary; some of these are shown in Fig. 1 along with explanatory captions. At 9 to 10 at. pct Sb, and thus above the solid-solubility limit Cs = 5.8 at. pct Sb and yet below the eutectic concentration Ceu = 17.5 at. pct Sb, there is a persistent anomaly in all the variations. Measurement was easy enough and standard enough so that we may forgo discussion of either method or possible error, in view of the forceful self-consistency shown by experiment. Solute distributions were measured along the bidirectional ingots in order to determine if a solute buildup had occurred at the CET. No anomalous changes in concentration were observed along the bars at either side of the solute anomaly of 10 at. pct Sb. The microstructures of the ingots were also examined across the range of composition investigated. At concentrations less than 9 at. pct Sb the micro- and macrostructures were typical of the cellular-dendritic growth usually reported for the columnar zone. Above the anomaly, both the columnar grain size and the amount of primary phase decreased. This microstruc-tural observation can be correlated with observations made of decanted interfaces. Below 9 at. pct Sb, the interface reveals the "normal" columnar dendritic structure; however above the anomalous concentration range, preferential growth of certain dendrites appears to form a "super dendrite" pattern. In Fig. 2 a microstructural view of the etched '(super dendrites'' shows their constitution both in terms of relative phase proportion and in terms of the ordinary smaller dendrites.

Jan 1, 1969

By F. W. Jessen, James N. Howell

INTRODUCTION The determination of cloud points has to date been limited to rather transparent oils, visual observation of the formation of a solid phase being the criterion for the standard ASTM D 97-47 method. The need for a quiclc and simple method for determining cloud points under dynamic conditions for opaque crude oils became necessary in a study of fundamental aspects of paraffin deposition from crude oils. PREVIOUS WORK The determination of cloud points in opaque oils by viscosity temperature curves is fairly new, Young' described a method employing the capillary tube principle. No information exists relative to the determination of cloud points of crude or refined oils under dynamic conditions. Shock' noticed that deposition of paraffin from crude oils took place at temperatures at or below the temperature at which a sharp inflection point occurred in the viscosity-temperature curve. This observation indicates a possible correlation between this temperature and the cloud point of the crude oil. Reistle and Vietti1 have reported that when crude oils containing small amounts of asphaltic materials were heated to temperatures sufficiently high to solubilize these materials and immediately cooled the pour point of the crude oil was lowered as much as 70°F. Sachanen\ asribes this effect to the action of colloidal asphaltenes in preventing the formation of the crystalline structure necessary to maintain the oil in a rigid state. At high temperatures, asphaltenes are peptized, and the paraffin crystals formed at these temperatures adsorb the colloidal asphaltenes on their surface thereby hindering further growth of the crystals. Reistle" found, however, that such heating and cooling of crude oils had no effect upon the cloud points. This latter work dealt with a crude oil of fixed composition, and the question immediately arises as to the effect of varying the concentration of asphaltic materials on the cloud point. APPARATUS AND PROCEDURE All viscosity measurements were made with a Bendix Ultra-Viscoson. The Bendix Ultra-Viscoson, which consists of a vibrating probe and an electronic computer. measures viscosity by exciting a thin alloy steel blade on the end of a probe by short electrical pulses in damped oscillation at 28 kilocycles. The blade vibrates longitudinally with an amplilude of less than 55 micron or 20 millionths of an inch. Energy to protlucc this excitation is generated in he computer and transferred to the probe through the connecting cable. The oscillating blade is magnetostric-tive, i.e., it is capable of transforming electrical energy to mechanical energy in the ultrasonic range. Ultrasonic waves are produced in the material surrounding the blade by the oscillating motion of the blade causing layers of the material to slip back and forth over one another at a known frequency. The electronic computer calculates the energy required to produce the sliding motion which is in turn proportional to the viscosity of the material. Actual values of viscosity times density (centi-poise X grams/cc) are indicated on the computer and a continuous D-C signal is provided which is fed into a type 153 Electronic Recorder (manufactured by Brown Instruments Div. of Minneapolis-Honeywell Regulator Co.) The liquid sample container was fabricated from a Lucite cylinder and jacketed by a Luicite water bath. Early experiments using glass and steel sample containers were not satisfactory. since the Ultra-Viscoson was apparently affected by the alterzltion of the magnetic fields surrounding the probe by stresses set LIP in the glass and steel as it cooled. The procedure for determining viscosity-temperature curves was relatively simple. The temperature ot the sample was raised sufficiently high to solubilize all paraffin and asphalts present. After heating, the oil sample was immediately transferred

Jan 1, 1957

By Lillian Erskine

While brass and other copper alloys have long been listed as offering health hazards to their workers, it is questionable if the metals involved are alone responsible for the trades' records of morbidity. The entire science of foundry sanitation is the product of the past two decades; and the best standards now operative would have been regarded as little short of revolutionary prior to 1910. The special handicap to the conservation of the health of the brass founder's employees has been the fact that, unlike the iron founder, his business could be profitably carried on in almost any small and unsanitary quarters. As a result, writers on occupational disease have accepted brass chills or brass founders' ague as an inevitable feature of the trade; and respiratory disabilities, as well as phthisis and disturbances of the nervous system, are also included as prevalent among those employed. While there can be no question as to the basic accuracy of past investigations, there is every reason to doubt the necessity for an undue percentage of ill health among brass molders and casters who are safeguarded by scientific ventilating standards and by improving personal habits of sobriety. There is, however, need for a campaign of education as to the most approved practices and equipment whereby the employer may safeguard the health of his workers, and thereby reduce his labor turnover and increase his productive efficiency. Broadly speaking, copper, tin, and aluminum carry negligible health hazards. The toxic qualities of lead, however, are a menace; especially if scrap lead is recovered or bought for foundry processes. The rapid volatilization of zinc renders its fumes and flocculent condensation the chief problems in conserving the health of the brass worker. The frequent presence in the foundry air of traces of arsenic compounds, lead

Jan 1, 1919

By Robert Mehl

METALLURGISTS are properly interested in papers dealing with subjects ranging from the theories of the metallic state to very practical details on the use of metals. A review of theoretical metallurgy naturally excludes papers concerned with the usefulness of metals and limits itself to those concerned purely with the science of metals. Even in this more limited field the reviewer is confronted with thousands of papers dealing with facts and theories of metallic behavior, from which he can select only a few hundred. The selection is necessarily highly arbitrary. Certain classes of papers definitely fall within the scope of this review. Thus, papers on the deformation of metals, on the structure of alloys, on recrystallization, to select only a few, are fully represented. Other subjects, which verge on what is generally accepted as the science of metals, such as the magnetic properties of metals, the less familiar electrical properties and theories, adsorption of gases by metal surfaces, and the like, are treated scantily. Particular attention is given to reviews of subjects that fall with in the groups chosen; in a summary such as this, which is largely in the nature of a progress report, capably written reviews are of importance, particularly since the monograph and text book literature is so meager in this field.

Jan 1, 1935

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By Bunting S. Crocker

EBBLE grinding at Lake Shore is not a temporary BlE wartime substitute. The tube milling plant, with a 1000 ton per day capacity, grinds a hard siliceous ore to 90 pct — 325 mesh. The plant, prior to using pebbles, was consuming 4.3 lb of 11/4-in. grinding balls per ton of ore, which amounted to 785 tons of balls per year. At September 1951 prices, $132.60 per ton, this steel cost amounted to $104,400 per year, or $0.285 per ton milled. By the Lake Shore method of substituting screened rock for this steel, all of this cost is saved. This is one of the major economies in Lake Shore mill practice. Regardless of the ultimate price of grinding balls, a change back to steel balls is not considered. The present pebble plant is more flexible than a steel ball plant and equally efficient. For example, if it is desired to change the size of grinding media, the pebble charge in any mill can be changed completely in 4 to 5 days, as against 76 days to change completely a 1-in. steel ball charge. To change pebble size it is necessary only to change two sets of screens and clean out the rock feed storage bin. This will take 8 to 10 days as against 3 to 4 months to clean out the customary supply of grinding balls kept on hand. Also it has not always been possible to purchase all desired sizes of balls at any price. In an analysis of savings effected by the use of the pebble mills, the flexibility of the Lake Shore grinding plant should be discussed, as it has a direct bearing on these savings. The plant has always used several units to handle tonnage rather than sending all the tonnage through a single unit. This principle may result in using small diameter mills, but no objection to that is seen. At no time has any advantage been found in cost per ton ground in large diameter mills. Both the capacity and the power of any mill varies as the diameter raised to the 2.6th power. Consequently large or small mills are equally efficient, and a plant, should be designed to use as many units or combination of units as is consistent with reasonable operating practice. Mills under 5 ft diam are harder to reline, etc. To define the case at Lake Shore, 2600 tons per day formerly were milled in seven 7x6-ft ball mills and twelve 5x16-ft (and two 6x16-ft) tube mills.* This gave an excellent test plant and an extremely efficient one. In this plant the ratio of ball mills to tube mills was 1:2. When the much cheaper pebble mills were substituted for the tube mills, this ratio was changed to one ball mill to four pebble mills to take the greatest possible advantage of the cheaper operating mill, i.e., the pebble mill. This flexibility without loss in efficiency has been an important item in the cost savings. It is interesting to note that the use of pebbles for fine grinding was proved first in the laboratory in a 12-in. ball mill. In fact, since 1934 all testing on — 8 mesh material has been done in this 12-in. mill. Scope of the Tests A paper on fine grinding at Lake Shore Mines was published in July 1940.' This paper covered 7 years of intensive research on fine grinding as well as sizing methods and equipment, plant scale grinding tests on 5x16-ft tube mills with and without grate discharges both with 1-in. and -in. balls, the use of laboratory mills to evaluate plant changes, and several reports on classifiers and classification. In the following July the addendum report2 was added in which the idea of series-circuit grinding was introduced, and the results of running five stages of tube mills and bowl classifiers were shown. Since 1940, the ball milling end of the plant has been altered extensively as a result of tests on the use of 3-in. rods in 7x6-ft mills and the use of the Tyler repulping screen with from 7 to 14 mesh screens. These tests are lengthy and may be covered in a separate report later. The scope of this report is confined to ore ground by rod milling and ball milling until it passed through an 8 mesh Tyler Ty-rod screen. The —8 mesh screen undersize then was pumped to a primary bowl classifier in open circuit and the sands from the bowl sent to the primary pebble mills, see Fig. 1. In the pebble mill circuit the ore is ground to 90 pct —325 mesh (24 pct + 28 microns). In studying the flowsheet, attention should be paid to the efficiency of the classification equipment used. The Tyler repulping screen is an efficient machine on the 8 to 10 mesh separation, and the bowl classifier is equally efficient at the 325 mesh separation. Efficient classification is a necessity for series-circuit stage grinding. The ore is hard siliceous porphyry, 60 pct SiO,, 80 pct insoluble. Its grindability at different meshes has been shown near the top of the list in F. C. Bond's grindability tests.' Lake Shore is not shown, but an adjacent mine with identical ore, Wright-Hargreaves, is. Reasons for the Changeover Since 1936 grinding balls have been rising steadily in price with no sign of stopping. For a mill that used 4.5 to 5.2 lb of grinding balls for every ton of ore ground this rise represented an alarming increase in grinding costs. In many cases the quality of the grinding balls fell off as the scrap steel became more difficult to obtain. The ratio of tube mills to ball mills increased with the use of the Tyler repulping screen in the ball mill circuit. Originally only 3.0 lb of balls per ton were used in the tube mills, and this

Jan 1, 1953

By C. J. Ramsburg

The construction and operation of by product coke-oven plants in America are essential to strong national defense and of the greatest importance to many widely diversified undertakings as well as to safety and health. Carbonization Process Coal used under a boiler or in a locomotive is consumed by combustion. Coal used in a by-product oven is distilled or carbonized, meaning that heat is applied to the coal, out of contact with air, causing a distillation that removes the volatile matter in the coal in the form of gas and vapor, leaving coke as a final product. From the gas and vapor distilled from the coal, many valuable substances can be secured, known as by-products. Hence the name "by-product oven." The simplest example of this principle can be shown by means of an ordinary clay pipe. If powdered coal is placed in the bowl of the pipe, the opening of the bowl is plugged with clay, and the bowl placed in a flame, the bowl will become hot, transferring the heat through the wall of the bowl to the coal within. When this coal attains a temperature of about 750°F., gas will begin to pour out of the pipe stem, and this gas may be lighted. A black liquid also will exude from the stem of the pipe, This is tar mixed with certain other products. What remains in the bowl after distillation is coke. A coke oven corresponds to the bowl of the pipe. A modern oven of the latest type is a refractory brick chamber approximately 42 ft. long and 14 ft. high, averaging 18 in. in width. The oven is filled with coal through closable charging holes in the oven top. Each end of the oven consists entirely of a self-sealing door, the removal of which permits the finished coke to be discharged by means of a power-operated pusher machine. The ovens are built in batteries, and the heating is effected by gas burned in vertical flues in the oven walls, each set of flues heating the walls of two adjacent ovens. Such ovens carbonize a charge of approximately 18 tons of coal in 18 hr., or have a daily throughput of 24 tons per oven, producing about 17 tons of coke. In times of need, the output may be increased by raising the temperature of flues, thus shortening the coking time. There is one great essential difference in principle between distillation in a clay pipe and in a coke oven. In the 18-in. oven, the distillation proceeds from two parallel walls held at a high temperature (over zooo°F.) and instead of all the coal being heated at once, the distillation is a progressive one creeping inward. The average travel of this heating is 35 in. per hour from each vertical wall, and as the volatile products are formed progressively they must push away from the interior toward the wall through the previously formed hot coke, then up the walls and Over the top of the coal charge, finally escaping through the outlet pipe. This results in a cracking of the components into many different and valuable substances not formed when the carbonizing temperature is low.

Jan 1, 1942

By C. J. Ramsburg

The construction and operation of by product coke-oven plants in America are essential to strong national defense and of the greatest importance to many widely diversified undertakings as well as to safety and health. Carbonization Process Coal used under a boiler or in a locomotive is consumed by combustion. Coal used in a by-product oven is distilled or carbonized, meaning that heat is applied to the coal, out of contact with air, causing a distillation that removes the volatile matter in the coal in the form of gas and vapor, leaving coke as a final product. From the gas and vapor distilled from the coal, many valuable substances can be secured, known as by-products. Hence the name "by-product oven." The simplest example of this principle can be shown by means of an ordinary clay pipe. If powdered coal is placed in the bowl of the pipe, the opening of the bowl is plugged with clay, and the bowl placed in a flame, the bowl will become hot, transferring the heat through the wall of the bowl to the coal within. When this coal attains a temperature of about 750°F., gas will begin to pour out of the pipe stem, and this gas may be lighted. A black liquid also will exude from the stem of the pipe, This is tar mixed with certain other products. What remains in the bowl after distillation is coke. A coke oven corresponds to the bowl of the pipe. A modern oven of the latest type is a refractory brick chamber approximately 42 ft. long and 14 ft. high, averaging 18 in. in width. The oven is filled with coal through closable charging holes in the oven top. Each end of the oven consists entirely of a self-sealing door, the removal of which permits the finished coke to be discharged by means of a power-operated pusher machine. The ovens are built in batteries, and the heating is effected by gas burned in vertical flues in the oven walls, each set of flues heating the walls of two adjacent ovens. Such ovens carbonize a charge of approximately 18 tons of coal in 18 hr., or have a daily throughput of 24 tons per oven, producing about 17 tons of coke. In times of need, the output may be increased by raising the temperature of flues, thus shortening the coking time. There is one great essential difference in principle between distillation in a clay pipe and in a coke oven. In the 18-in. oven, the distillation proceeds from two parallel walls held at a high temperature (over zooo°F.) and instead of all the coal being heated at once, the distillation is a progressive one creeping inward. The average travel of this heating is 35 in. per hour from each vertical wall, and as the volatile products are formed progressively they must push away from the interior toward the wall through the previously formed hot coke, then up the walls and Over the top of the coal charge, finally escaping through the outlet pipe. This results in a cracking of the components into many different and valuable substances not formed when the carbonizing temperature is low.

Jan 1, 1942

By B. T. Wykoff

This paper describes an important function of mining in Southeast Missouri. The practice is a necessary procedure to prevent disintegration of the pillars. Although the technique is peculiar to this district, it might have application in other mines using room and pillar method. CONSIDERING the subject alone, this paper would cover only a mechanical treatise of pillar roping. Because the St. Joseph Lead Co. originated the practice, it is proper to give a short introduction to the company. It is also proper to describe pillars and discuss their purpose to justify expending large sums of money to preserve them after the surrounding area has been thoroughly mined. Because the St. Joseph Lead Co. originated the practice of back pinning with wedge bolts, patch plates and channel irons, and the pouring of substitute pillars of concrete, these too, are mentioned. The Southeast Missouri Division of the St. Joseph Lead Co. is the largest lead producer in the United States. About 95 pct of the present producing mines are in St. Francois County, which begins 60 miles south of St. Louis. The entire lead ore production of this county, at the present time, is mined and milled by this company, treatment plant capacity being about 22,000 tons per day. This area is also the oldest, lead having been mined at Mine La Motte, in Madison County, in 1723. Estimated production of all the various companies that have mined here is about 7,000,000 short tons of pig lead. Method of Mining The lead occurs in the ore as galena. Galena is found either in the form of disseminated or solid streaks, of varying thickness. Ore stratum is the Bonne Terre dolomite, which is about 350 ft thick, but almost all the ore is in the lower 100 ft. Ore is mined by room and pillar method in open stopes. These stopes vary in height from 7 ft to 200 ft, depending on the thickness of the pay ore in that spot. Back, or roof, is supported entirely by pillars. In a majority of stopes, after the loose scales are mined down, the back has remained solid for many years (fig. 1). Under ideal conditions pillars are spaced to form an equilateral triangle and are as small and as far apart as the height of the stope and the character of the back will permit. They vary in diameter from 10 ft to 50 ft and are from 18 ft to 50 ft apart, measured from the outer edges. In general, the pillars are approximately 12 ft in diameter, and 25 ft apart. The spacing is very uncertain because of the vagaries of the ore trend, the desire to better support breaks in the back, and the attempt to spot out the thinner ore by not placing one in the core of the ore body. Use of Structural Steel Channels In stopes where the back is bad because the overlying formations are thin with wet, weak and shaly

Jan 1, 1951

By B. T. Wykoff

This paper describes an important function of mining in Southeast Missouri. The practice is a necessary procedure to prevent disintegration of the pillars. Although the technique is peculiar to this district, it might have application in other mines using room and pillar method. CONSIDERING the subject alone, this paper would cover only a mechanical treatise of pillar roping. Because the St. Joseph Lead Co. originated the practice, it is proper to give a short introduction to the company. It is also proper to describe pillars and discuss their purpose to justify expending large sums of money to preserve them after the surrounding area has been thoroughly mined. Because the St. Joseph Lead Co. originated the practice of back pinning with wedge bolts, patch plates and channel irons, and the pouring of substitute pillars of concrete, these too, are mentioned. The Southeast Missouri Division of the St. Joseph Lead Co. is the largest lead producer in the United States. About 95 pct of the present producing mines are in St. Francois County, which begins 60 miles south of St. Louis. The entire lead ore production of this county, at the present time, is mined and milled by this company, treatment plant capacity being about 22,000 tons per day. This area is also the oldest, lead having been mined at Mine La Motte, in Madison County, in 1723. Estimated production of all the various companies that have mined here is about 7,000,000 short tons of pig lead. Method of Mining The lead occurs in the ore as galena. Galena is found either in the form of disseminated or solid streaks, of varying thickness. Ore stratum is the Bonne Terre dolomite, which is about 350 ft thick, but almost all the ore is in the lower 100 ft. Ore is mined by room and pillar method in open stopes. These stopes vary in height from 7 ft to 200 ft, depending on the thickness of the pay ore in that spot. Back, or roof, is supported entirely by pillars. In a majority of stopes, after the loose scales are mined down, the back has remained solid for many years (fig. 1). Under ideal conditions pillars are spaced to form an equilateral triangle and are as small and as far apart as the height of the stope and the character of the back will permit. They vary in diameter from 10 ft to 50 ft and are from 18 ft to 50 ft apart, measured from the outer edges. In general, the pillars are approximately 12 ft in diameter, and 25 ft apart. The spacing is very uncertain because of the vagaries of the ore trend, the desire to better support breaks in the back, and the attempt to spot out the thinner ore by not placing one in the core of the ore body. Use of Structural Steel Channels In stopes where the back is bad because the overlying formations are thin with wet, weak and shaly

Jan 1, 1951

By M. G. Nicholas

The influence of deformation, relative movement and surface cleanliness on the strength of room -temperature "interparticle" bonds between copper surfaces has been studied. No bonding resulted when clean surfaces were merely pressed together, even under pressures of 50,000 psi. Bonding resulted, however, when clean surfaces were simultaneously pressed and moved relative to each other, some bonding occurring no matter how small the pressure or relative movement. Exposure of the clean surfaces to various chemicals inhibited bonding, none at all being produced by small relative movements. The size of the movements which first produce bonding between exposed surfaces depends on the nature of the chemical, being less than 0.002 in. for cleansing reagents and greater than 0.004 in. for lubricants, water and hydrogen sulfide. It is concluded from the experimental data that the prime requirement for bonding is the dispersal of surface films such as oxides and adsorbed contaminants. THE formation of interparticle bonds during the room-temperature compaction of powder metallurgy components greatly increases the ease of handling in subsequent production operations prior to sintering. The strength of unsintered compacts follows the same general trends as their density and hardness, but quantitative data are extremely Scarce. Quantitative correlation of the data on the overall strength of a compact with that of the individual interparticle bonds is virtually impossible due to the extremely complex geometry of powder compacts. Qualitatively, however, the strengths of unsintered compacts can be correlated with variations in parameters known to influence the extent of cold bonding, and hence interparticle bonding, such as surface cleanliness, relative movement, and deformation of the mating surfaces. Kuzmic and seelig4 found that contamination of powder particle surfaces with lubricants reduced the strength of unsintered iron and copper compacts. Interparticle shear has been shown to promote bonding by Train and Hersey7 while Gregory has shown that die rotation is beneficial. Eilender and schwalbe6 concluded that most bonding occurred during the early stages of pressing iron powder compacts when interparticle shear is one of the principal processes occurring. mile there can be little doubt that these and other factors, such as particle size and shape, affect the extent of interparticle bonding produced by compaction, evaluation of the influence of individual factors is not easy. Particle deformation and relative movement, for example, occur simultaneously during compaction, making any precise measurement of their relative influences difficult. Because of these and other difficulties encountered in studying interparticle bonding in actual powder compacts, the approach adopted in this study was to

Jan 1, 1963

By G. Caetani, E. Burt

Jan 1, 1907

By J. H. Nead

OBJECT THE experiments herein described were undertaken with a view to investigating thoroughly the influence of carbon on the tensile and impact physical properties of carbon steel. The original comprehensive plan included investigation of a series of steels with varying carbon contents over a wide range of heat treatments. These were to be studied as to their static tensile qualities, and also their resistance to shock as measured by the Charpy impact bending test. A thorough metallographic examination was likewise included in the schedule. The Charpy tests and the metallographic examination have not been completed to date, and are accordingly withheld for another report. All heat treatments, however, have been made, and the static tensile tests have been completed. A description of the procedure of the experiment and the results thus far obtained are therefore reported herein. CONCLUSIONS No attempt is made to draw general conclusions from these experiments, but the curves and the data are submitted as being of general importance and wide applicability.

Jan 12, 1915

By G. K. Manning, M. C. Udy, L. W. Eastwood

Those who have examined beryllium and beryllium-rich alloys under the microscope have noted the results of the difficulties encountered when preparing these materials for examination. Hard constituents are readily chipped and pulled out of the matrix, and the soft ones are easily gouged out or embedded with abrasive or other material. The matrix is easily deformed, which makes it very difficult to remove the effects of scratches. Furthermore, the matrix and some of the constituents are also readily pitted during etching, which makes the structure difficult to develop. The metallographic techniques designed to avoid these difficulties are neither radically new nor necessarily the best possible procedures. They were developed at Battelle to fill immediate requirements as part of a study of the preparation and pouring of beryllium melts as described in a separate paper. A part of the research program on the causes of gas unsoundness in beryllium castings entailed a study of the effects of such gases or gas formers as oxygen or oxides, nitrogen or nitrides, and carbon or carbides. These gases or gas formers, if dissolved in the melt, might be an important factor in the study of unsoundness caused by gas evolution in beryllium. Because it was necessary to be able to distinguish these gases or gas formers, if present as alloy constituents, from metallic phases also present as alloy constituents, a series of alloys was prepared with additions of the various possible metallic and non-metallic constituent formers. It should be emphasized that the constituents referred to as gases or gas formers are important in the study of gases in beryllium only if they form a part of the alloy, that is, if they dissolve in the liquid or solid metal. During the first part of the development of metallographic techniques, the Massachusetts Institute of Technology, unpublished report (TC 3315, Nov., 1945) by Paul Gordon, describing various microconstituents in beryllium, was quite useful. Of particular interest in this report is the confirmation of one of the conclusions drawn from the present investigation, that is, that a distinct oxide phase apparently does not occur in beryllium. The Preparation of Metallographic Specimens GRINDING PROCEDURE The original specimen is either sawed or cut on an abrasive wheel to a convenient size. For best results, the surface to be polished should not be more than about 1/4 in. square, and it is convenient to mount the specimens in bakelite to facilitate handling. Two alternative grinding procedures have been developed. As one alternative, specimens are ground successively on 120-, 240-, and 400-grit, wet-or-dry metallographic discs. The abrasive discs are revolved at 1750 rpm on a conventional pedestal grinder. The coarsest disc (120 grit) is used either wet or dry, and in most instances, it can be eliminated from the procedure. Grinding on the two finer discs (240 and 400 grit) is accompanied by the application of kerosene. The technique consists of holding an oil can, containing kerosene, in one hand and the mounted specimen in the other hand. While the grinding is being done, several drops of kerosene are applied every few seconds close to the center of the disc. Carbon tetrachloride serves at least as well as kerosene as a lubricant; perhaps it is slightly better, but it volatilizes so readily that an additional health hazard is involved. Light oils and water are not satisfactory for the finer grinding operations. The second procedure is perhaps slightly slower than the first and requires a more careful technique, but it eliminates the use of kerosene. The specimen is first ground wet on a 240-grit disc, followed by dry grinding on a 400-grit disc. Finer grinding does not appear to be necessary with either of the two alternative procedures. The pressures used throughout either grinding operation must be extremely light, that is, barely sufficient to hold the specimen in contact with the disc; otherwise, flow and chipping are almost certain to occur. In both procedures, it is essential that the discs be sharp and that they be discarded as they show evidence of becoming dull or loaded. In general, not more than four or five specimens can be ground on a disc before the disc becomes dull enough to warrant discarding it.

Jan 1, 1950

By M. C. Flemings, P. J. Ahearn

Dendvite morphology of unidirectionally solidified Sn-12 pct Bi alloy was determined with the aid of X-ray and macroetch techniques. Columnar growth direction (growth direction of primary arms) in this alloy is [110]; secondary arms grow in the [lii], [111), and [112] directions. Interstices between [112] dendrite arms fill in during early stages of solidification so that well-defined (110) planes are seen in the final structure (parallel to the heat-flozo direction). Similarly, interslices betweex the [111] secondary arrlls fill in to form (112) planes, and interstices between [1111 ] arms fill in to form (112) planes. DENDRITE morphology in alloys is conveniently studied in unidirectionally solidified castings. In this type of solidification, heat is extracted from a single chill surface of the casting while other surfaces are insulated. Solidification progresses from the chill and results in cast grains that are columnar with the long axis of the grain perpendicular to the chill. Dendrite morphology is conveniently studied by examination of etched surfaces parallel and perpendicular to the chill, as previously described.'" Experimental studies on the growth direction for tin are in general agreement. The major growth direction is [110]:-' although this is altered somewhat by supercooling. In one study, tin dendrites pulled from a supercooled melt had a growth orientation 12 deg away from the [110] direction toward the [002] direction.' This orientation could be changed back to the [110] or close to the [110] direction by changing the tempera- ture gradient in the melt. In a study using autoradio-graphic techniques, (111) directions were shown to be directions of preferred growth in addition to the [110) direction.7 PROCEDURE The samples employed in this study were cut from 12 pct Bi-88 pct Sn castings, 1.5 in. in diam and 3.5 in. tall. The casting procedure consisted of heating the charge to 390°C and then pouring the molten alloy directly into a gypsum sleeve mounted on a massive copper chill. Specimens were cut from the ingot, mounted, polished, and etched with Taffs reagent [FeC13 .6H2O, 2 g; HC1, 5 cu cm;H20, 30 cu cm; CzHaOH (95 pct), 60 cu cm].' The dendritic pattern was clearly visible after etching. Photomicrographs at about 8 and 35 times were taken on surfaces parallel and perpendicular to the chill. Laue back-reflection patterns were obtained from three different columnar grains in a singIe polished face 2 in. from the chill. The cut face was polished and positioned in the beam of a Laue camera, so that diffraction occurred from a single dendritic grain. The resulting diffraction pattern was interpreted to determine crystal log raphic orientation in relation to the heat-flow direction and in relation to microstruc-ture of the given grain. RESULTS Fig. 1 shows a horizontal and vertical section of the columnar structure, at low magnification. Light regions are rich in tin, and darker regions tin-poor; the primary phase is tin-rich a phase. The two sections shown were taken orthogonally from the same metallographic specimen. It is readily seen from Fig. 1 that the dendrites have

Jan 1, 1968

By Charles Fettke

IN 1914, the officials of the Brundred Oil Corpn., faced with the problem of introducing new methods to increase production in the old and nearly depleted pools of Venango County, became interested in the Smith-Dunn (Marietta) process which was at that time being tried out in southeastern Ohio and which was already yielding very encouraging results. A license to use the process in Pennsylvania was secured from the owners of the patents. These patents were later declared invalid. The method was first used in Pennsylvania on a property consisting of about 434 acres, of which 300 are productive, at Hamilton Corners, 15 miles northeast of Oil City, in Cherry Tree and Oakland Townships, Venango County. The producing horizon is the First Sand of the Venango Group. A small oil pool was opened up in this vicinity in 1904. By July, 1916, when the compressed air was first introduced to restore pressure, the production had dropped to a point where it was no longer possible to continue profitable operation. From that time until March, 1927, over 100,000 bbl. of oil have been recovered from the tract. The property is being operated at present by the Brundred Oil Corpn. and Lux. The 1926 production, amounting to 10,650 bbl., was only 400 bbl. short of the maximum reached during any year of the 10-year period that the process has been in operation. Inasmuch as the Hamilton Corners pool is the first one in Pennsylvania in which the Marietta process was tried out, and the one in which it has been longest in continuous operation, considerable interest attaches to the results obtained. Through the courtesy of H. D. Brown and W. J: Brundred, officials of the Brundred Oil Corpn., all their data on this operation were made available to the writer in connection with an investigation of rejuvenation methods in the State for the Pennsylvania Geological Survey.

Jan 1, 1928

By Henry Lewelling

An experimental investigation of the relative effectiveness of standard bullets and "shaped charges" in perforating dense, hard formations is reported. A method is described which simulates the conditions under which a shot is normally fired. Results are evaluated on the basis of depth of penetration, extent 01 Fracturing. extent and nature of formation damage, and other physical characteristics. INTRODUCTION A study was made of the relative merits of bullets and "shaped charges'' as applied to well completions in hard, dense formations such as those encountered in the Permian Basin. The planning of the laboratory program stipulated that core samples should be prepared and mounted in such a way that perforation shots of either kind"" could be fired into them, keeping the factors pertinent to good shooting practice similar to what they would be in a well. The shooting was done in a buried tank. with the sample and shooting equipment covered with water. The depth of water was calculated to he sufficient to simulate the inertial effects of well fluids for the duration of an explosion (a few micro-seconds). Flow measurements were attempted on the specimens after perforation. Due to the lack of uniformity of the limestone, shale breaks. etc., these measurements proved to be of little value to the study. The mounting materials were removed. so that the character of the penetration. fracturing, etc., could be observed. METHOD Laboratory samples were selected from the hard. dolomitic sand cores from Block 31 Field. Core pieces having a 3 3/4 in. diameter and a useful length of 10 to 12 in. were used. Selection was aimed at those pieces having a minimum of shale streaks and other irregularities. The ends were cut off square Manuscript received in the Petroleum Branch office June 28, 1951. Paper presented at the Petroleum Branch Fall Meeting in Oklahoma City, Okla., Oct. 3-5, 1951. with the diamond saw. Prior to lbeing mounted. the cores were soaked in water to prevent them from absorbing water out of the wet concrete, thus weakening it. A set of steel capsules was fabricated for holding the core Samples. Seven-in. casing was cut in two-ft lengths. of which one end was closed by "orange-peeling" and welding, and the other was threaded to accept a standard bull plug. The bull plugs were fitted with copper tube connections .so that pressure could be applied when the capsules were made up. The cores were mounted in the capsules by means of wire centralizers whose function was to hold the Core in proper position while concrete was poured and set. A special aggregate was prepared wing a graded sand and gravel mixture

Jan 1, 1952

By Kennedy, Joseph E.

MANY things must be considered in coming to a decision as to what type of crushing and grinding equipment is to be used for preparing run-of-mine ore for concentration, amalgamation, flotation, or cyanidation. This paper is intended in no sense as a thorough discussion of these various factors, but rather contains a few observations based on a wide experience in the manufacture and application of milling machinery. One important governing factor is the tonnage to be dealt with. In large plants treating several thousands of tons in eight hours, a large jaw or gyratory crusher of substantial construction is standard equipment for reducing ore to 8 or 10 in. or finer. In open-pit work, using power shovels with 4- or 5-yd. dippers, a jaw crusher should have a receiving opening of 48 by 60 in., or 60 by 84 in., so that dynamite or hand sledging will not be required for breaking to a finer size before primary crushing. The coarsely broken ore is then conveyed by a belt to vibrating screens, the oversize from which is again conveyed by belts to secondary crushers having openings of 12 or 14 in., the product from which is conveyed to storage. From storage, the material goes to a finishing gyratory for reduction to 1/4 in. or finer.

Jan 1, 1936

By John W. Buch

IN my review a year ago I pointed out that a small coal-mining companies as well as large had decided that the so called ?central shop? was a benefit. These central shops replaced in a large measure the commercial shops operated by those who have no direct connection with production of cod. Central shops in this time of salvage and major repairs have proved their worth. In addition, many of the larger companies had in their central shops equipment available for the production of many thousands of parts for our vast war machine and, as subcontractors, are doing this type of work. This is a real contribution to the war endeavor. Lack of new supplies has resulted in exhaustion of those stocks formerly stored or hidden in underground shanties. This means that the small stock of repair parts likely LO be available to any mine

Jan 1, 1943