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By J. H. Slater

There is nothing particularly new about the use of sinter in a blast-furnace burden. For many years flue dust has been sintered at the various blast-furnace plants to put it in a form that could be recharged to the furnace with some assurance that it would stay in the second time. This paper is not a compilation of scientific data but only an attempt to give our experience with sinter. There are some unexplained discrepancies in the tests given here, such as the flue dust produced and iron tonnages, but the results given are the results that were obtained. Sinter is not the cure-all for blast-furnace problems. A good coke is still the most important factor in good furnace operation. However, good sinter is a big help. As flue dust is produced under varying operating conditions and fluctuates greatly in carbon content, it is difficult to produce from flue dust alone a uniform sinter, which will give satisfactory blast-furnace results. Sinter made from flue dust alone will change in character with every carbon variation. Almost invariably flue dust contains carbon in excess of the amount necessary to make good sinter, and this excess must be burned out. This, of course, necessitates reduction of the speed of the sintering machine and thus a decrease in tonnage. An alternative is to dilute with a carbon-free material to use up the excess carbon. High-carbon material when sintered results in a very dense, undesirable product, since the excess heat melts the charge into a hard, nonporous mass, and tends to form refractory silicates. At the Republic Steel plant in Cleveland, there had existed since 1924 a continuous-type sintering machine having 224 sq. in. of grate-bar area, capable of producing 9 to 11 tons of sinter per hour from flue dust. Under a former management, it was the practice to sinter only the flue dust from the blast furnaces, no attempt being made to recover the gas-washer sludge. As the dust was limited, the plant operated only part time. In other words, there was a sintering plant representing considerable investment, with good facilities for screening and preparation of raw materials, and flue dust available for only part-time operation. The sinter produced was costly and not a very desirable product. The first step was to install equipment for the recovery of the gas-washer sludge. This was done in the form of two large thickeners and a filter, which have proved very satisfactory. It was then decided to screen out the minus 3/8-in. fines from the Mesabi ores, mix them with flue dust and coke breeze when necessary, and operate the plant at full capacity. The tonnage increased to 18 tons per hour and the product was much improved. Sintering costs fell and the product was a preferred blast-furnace material, which on an iron unit cost compared very favorably with good Mesabi ores. On the average, 40 per cent flue dust and 60 per cent fine ore was being used. The oversize from the ore screening also became a preferred product for the

Jan 1, 1941

By J. H. Slater

There is nothing particularly new about the use of sinter in a blast-furnace burden. For many years flue dust has been sintered at the various blast-furnace plants to put it in a form that could be recharged to the furnace with some assurance that it would stay in the second time. This paper is not a compilation of scientific data but only an attempt to give our experience with sinter. There are some unexplained discrepancies in the tests given here, such as the flue dust produced and iron tonnages, but the results given are the results that were obtained. Sinter is not the cure-all for blast-furnace problems. A good coke is still the most important factor in good furnace operation. However, good sinter is a big help. As flue dust is produced under varying operating conditions and fluctuates greatly in carbon content, it is difficult to produce from flue dust alone a uniform sinter, which will give satisfactory blast-furnace results. Sinter made from flue dust alone will change in character with every carbon variation. Almost invariably flue dust contains carbon in excess of the amount necessary to make good sinter, and this excess must be burned out. This, of course, necessitates reduction of the speed of the sintering machine and thus a decrease in tonnage. An alternative is to dilute with a carbon-free material to use up the excess carbon. High-carbon material when sintered results in a very dense, undesirable product, since the excess heat melts the charge into a hard, nonporous mass, and tends to form refractory silicates. At the Republic Steel plant in Cleveland, there had existed since 1924 a continuous-type sintering machine having 224 sq. in. of grate-bar area, capable of producing 9 to 11 tons of sinter per hour from flue dust. Under a former management, it was the practice to sinter only the flue dust from the blast furnaces, no attempt being made to recover the gas-washer sludge. As the dust was limited, the plant operated only part time. In other words, there was a sintering plant representing considerable investment, with good facilities for screening and preparation of raw materials, and flue dust available for only part-time operation. The sinter produced was costly and not a very desirable product. The first step was to install equipment for the recovery of the gas-washer sludge. This was done in the form of two large thickeners and a filter, which have proved very satisfactory. It was then decided to screen out the minus 3/8-in. fines from the Mesabi ores, mix them with flue dust and coke breeze when necessary, and operate the plant at full capacity. The tonnage increased to 18 tons per hour and the product was much improved. Sintering costs fell and the product was a preferred blast-furnace material, which on an iron unit cost compared very favorably with good Mesabi ores. On the average, 40 per cent flue dust and 60 per cent fine ore was being used. The oversize from the ore screening also became a preferred product for the

Jan 1, 1941

By H. M. Zoerb

Based on the belief that operating details are a definite contributing factor to major economies, this paper traces the development of crushing cavity design in Symons cone crushers to attain maximum liner utilization. Wear rates are analyzed and compared in this presentation and drawings illustrate succeeding design changes. IN these times of rising labor and material costs, it has become more and more necessary that attention be paid to some operating details which, in their obscurity, may he the key to major economies. Liner wear in crushing cavities of secondary and tertiary crushers can become an appreciable cost item when the material to be crushed is hard and abrasive. This item of cost not only includes the value of the crushing members, but also more intangible costs such as labor and lost production due to more frequent replacement. The variables which are encountered in ores and minerals to be reduced; the design of plant and machine application; the sizes, shape, and fineness, characteristics of the crushed product; the moisture; hardness; friability; and abrasiveness of the material to be crushed are all influencing factors which must be taken into consideration in the selection of a crusher, and particularly in the design of crushing cavity and liners to be used in a crusher. Through a research program undertaken in cooperation with many operators of Symons cone crushers a new approach to crusher cavity design was made, resulting in the development of liners for specific operations which showed: 1—maximum utilization, as high as 70 to 80 pct of original weight of metal, and 2— maximum capacity of unit during the greater portion of its life. It has been found that liners so designed for a given operation will show added economies in power consumption, maintenance, and general wear and tear on the crushing unit. Initial work in the so-called tailoring of crushing cavitles was begun on the tertiary or fine crushing units where as a rule reduction ratios were low, varying from 3 to 6. Parallel or sizing zones in the lower portion of the crushing cavity were too long, resulting in a tendency to pack. It was found that very little additional crushing was done in the parallel zone after the initial impact in that zone and that a relatively small amount of' additional crushing was done by attrition, which required very careful feed control. A small amount of over-feeding would result in packing which not only consumed power but caused unnecessary liner wear as well. The illustrations which follow in this discussion will show only contours of crushing cavities, and for purposes of simplification the cavities will be considered only in their closed position. The first step, therefore. was to reduce the sizing zone to a minimum. This was done by removing the lower portion of the liner as shown in Fig. 1. The result of the change was a saving of 15 to 20 pct in liner cost, less power consumption, with no change in capacity. This change in design, while an improvement, did not go far enough. As wear took place, the change in the liner was not uniform throughout its entire length, resulting in a restriction of the feed opening and thereby loss of capacity. Furthermore, progressive wear of the liner had the effect of lengthening the parallel zone until finally the entire crushing cavity was all parallel zone, see Fig. 2. It is obvious from the reduced feed opening of the worn liner that the ability of the machine to receive material is lessened considerably. Furthermore, the long parallel zone with its worn, irregular profile did not operate at its highest efficiency. The first attempt to overcome this difficulty was carried out on a 5 1/2-ft crusher installed in a plant producing roofing granules. The material being crushed was a very hard graywacke and the crusher was closed-circuited with a screen having .232-in. slotted openings. A radical change in contour was developed, as illustrated in Fig. 3. Equal wear lines on both concave and mantle are designated 1, 2, 3, etc. The method of development of this contour is as follows: Since adjustment for wear is vertical, corresponding intersections of wear lines and vertical lines developed concave and mantle contours which maintained equal but lengthening wear surfaces in the parallel zone. The ideal contour, of course, is one in which the length of the parallel zone remains constant, but because of present foundry practice and heat treating characteristics this is impossible.

Jan 1, 1954

The thirty-ninth meeting of the Boston Section was held at the Engi- neers? Club on Monday evening, Dec. 4, 1916. Thirty-three members and guests sat down to dinner; There were present as section guests; Dr. L. D. Ricketts, and Prof. Bradley Stoughton, President and Secretary of the parent society, and each was called upon to speak. Mr. Stoughton spoke of the growth of the Institute during the past year and described the very numerous and important activities, of which the junction of the American Society of Civil Engineers with the other three societies is perhaps the most important. Dr. Ricketts, the advertised speaker of the evening, spoke on Recent Progress in Mining and Metallurgy of Copper Ores. In discussing the improvements in mining, he traced the mining methods from square- set mining to the modified Ohio System now so successfully used at Inspiration: In metallurgy, he described the growth of reverberatory-smelting ant1 allied problems such as the utilization of the waste heat for generating steam. He laid stress on the importance of sampling and bedding of

Jan 6, 1917

By James Barr

Geology and Mineralogy1 TENNESSEE phosphates are commercially divided into three varieties: Brown, blue and white. The first two only are now of commercial importance. The white phosphates of Perry county are not being mined at present owing to their pockety occurrence. The white phosphate is Fig. I.-DEPOSITS OF 75 PER CENT. BROWN PLATE ROCK UNDERLAIN BY SAND. the highest grade of the three varieties, running often more than 90 per cent. tricalcium phosphate. The brown rock is found in three varieties, grading from the dense almost vitreous gray rock, 76 to 82 per cent. tricalcium phosphate, down

Jan 9, 1914

By Albert F. J. Bordeaux

Alluvial gold was first discovered in Guiana in 1852, in the sands of the Arataye river, by Paulino, a Braziliaii convict. During the following years, gold was found also in the rivers Orapu, CirubQ, etc. In 1873 the famous placers of the Sinnamary system were opened: Saint-Elie, Dieu-Merci, Adieu-Vat, and Couriege. This group of mines has produced nearly $10,000,000, according to official records, not including the gold that escaped the taxes. The discoverer was a mulatto named Vitalo. Except in a few rich spots, the gold was rather evenly distributed along the creeks. In one instance, gold was found in the red dirt from the slope of a hill, which, after transport to the creek, gave an amount of 150 kilograms. The placers of the Mana system, Enfin, Elysee, Pas-Trop-Tot, date from 1878, and their production has been about %g5,000,000. In 1888, the Awa placers, along the boundary bet.ween the French and Dutch Guianas, on the Maroni and its tributaries, were discovered. They have yielded $12,000,000. The only industrial company, the Dutch Guiana Freiich Minillg Co., produced in 15 years about 2,500 kg. of gold, worth $1,500,000. From 1893 dates the famous Carsewene discovery, in territory then in dispute between France and Brazil. A settlement was made recently in favor of Brazil, by arbitration of the Switzerland State Council. Carsewene is said to have produced about $18,000,000 in six years (1894-1900), but not even one-fourth of tho production went through French Quiana and

Jan 1, 1911

By Warnie Flint

According to statistics that have been compiled by the National Safety Council, U.S. Bureau of Mines, American Medical Association, and other agencies, accidental injuries cause more deaths than all the leading diseases. There isn't anyone who is immune to an accident. Probably the best way to describe an accident is to say that it is an unexpected happening. The automobile driver isn't expecting a terrifying crash, the electrician isn't expected to be electrocuted, nor is the coal miner expecting the roof to fall on him. All personal injuries will not result in death. There is a multitude of victims who are living proof of what accidents can mean to an individual. These are the victims who are blind, paralytic, missing arms, legs, etc. In most instances the earning power of these individuals is gone, and they are utterly dependent upon family or welfare. For many years there seemed to be an attitude that prevailed upon industrial America that accidents were inevitable, and that so many fatalities and injuries should be expected. This, of course, was utter nonsense. Since the birth of such organizations as The National Safety Council, American Society of Safety Engineers, and the U.S. Bureau of Mines, much progress has been made in reducing accidents. The enactment of compensation laws greatly contributed to the progress that has been made. Does this mean that all accidents can be prevented? Unfortunately, it doesn't. Human psychology indicates that men will work unsafely or perform unsafe acts for the following reasons: 1) Lack of knowledge or skill. 2) Conflicting motivations, such as: (a) to be more comfortable, (b) to save time or effort, (c) to be accepted by a group 3) Physical or mental condition.

Jan 1, 1973

By H. A. Wentworth

The Huff electrostatic plant of the United States Smelting CO., operated in conjunction with its wet concentrator at Midvale, Utah, was the second plant of substantial size installed using the Huff process, and the first plant to be put in operation on the so-called Western complex ores. , In spite of the machinery being of early design and consequently embracing many of the mechanical weaknesses incident to a pioneer plant the mill has operated steadily and uniformly since the summer of 1909, about five years now, without any material change other than some rearrangement of the machinery for better handling of the sizes. The ore, at present coming almost entirely from the company's mines in Bingham, is of practically the same composition as that upon which the plant was first put in operation. The crude ore, consisting of the sulphides of copper, lead, zinc, and iron, and containing small amounts of gold and silver, is brought by train and delivered to the hoppers of the wet concentrator, where by jig and table treatment, there are produced a shipping lead concentrate, a tailing, and a middling product, the latter being conveyed in push cars to the adjoining electrostatic mill. The crude ore delivered to the wet concentrator contains from 6 to 9.5 per cent. zinc. At times the plant is used also for the treatment of custom ores from the Bingham district and elsewhere, and at such times the results of the plant vary according to the material in use, but as by far the larger portion of the product is that from the company's own mines, the results given below are fairly indicative of the general work obtained. The diagram, Fig. 1, illustrates graphically the flow of the ore through the electrostatic mill. The middling coming from the wet mill, containing from 15 to 18 per cent. moisture, is hoisted while on the cars by an elevator, and delivered from the top of the elevator to a hopper over the drier. The drier first installed was of too low capacity to take care of the moisture present in the tonnage to which the mill was later

Jan 1, 1915

By R. P. Heuer

In recent years an increasing amount of research work has been done on refractory materials for use in copper-smelting furnaces. A few of the larger refractories manufacturers have erected special research laboratories which are fully equipped to study the refractory problems of the smelter. A typical laboratory of this type was described in the January, 1929, issue of Mining and Metallurgy, in an article entitled "Methods of Research Newly Applied to Refractories," by W. F. Boericke. This particular laboratory is equipped for the manufacture of full-size refractory brick. It aims to study on a fairly large scale the behavior of various refractory raw materials and the effect of improved methods of manufacture. The refractories manufacturer has at his disposal the naturally occurring refractory oxides in various degrees of purity, mixed or combined with other oxides. In order to supply the copper smelter with a refractory brick that will permit the most economical operation of the smelting furnaces, it is necessary to have the optimum combination of a number of both physical and chemical properties in the brick. Unfortunately, nature has not provided any raw materials in which the perfect combination of these properties exists. It is, therefore, a problem for man's ingenuity to seek out the most desirable raw materials and develop methods of manufacture that give, in the resultant product, the necessary combination of properties in as full a measure as is possible. In the laboratory it is possible to carry out fairly broad experiments with this objective. It is first necessary to obtain and process the raw materials, then to manufacture them into brick and finally, to test these bricks under conditions designed to interpret or simulate in the laboratory the actual conditions in the smelting furnace. It is necessary for the research engineers to determine certain fundamental properties in the brick which may be subjected to laboratory measurements and to interpret these measurements in terms of brick performance in smelting-plant practice. Such procedure is not what might be called an exact, scientific one but it has been found that great progress can be made when these principles are applied. For example, the resistance of refractory brick to spalling and thermal change can be related to the thermal expansion of the refractory, its thermal conductivity and heat capacity and its

Jan 1, 1934

By C. W. Wark, Fred Wise

The Getchell mine, a comparatively recent gold discovery, is in the old Potosi mining district, Humboldt County, Nevada. All ore is mined from open pits using Diesel shovels and gasoline trucks. Two types of ore are mined, oxide and sulphide. Oxide ores are siliceous or shaly; sulphide ores either gougy, siliceous or shaly. In mining, the sulphide ore is segregated into hard and gougy ore, because the gouge must be dried before it can be crushed. The oxide ore is not segregated. General Metallurgy The oxide ore is the leached outcrop that overlies the sulphide ore body. All of the iron and arsenic sulphides found in the sulphide ores have been thoroughly oxidized. The sulphur and arsenic have been leached out while the iron is present as hematite and limonite. The oxide ore is amenable to direct cyanidation, and although no visible free gold is present it is released at a comparatively coarse grind. The sulphide ore is more refractory, since the gold is all of micron size and some of it is locked in fine sulphides. It has been established microscopically that the free gold is coated with an impervious shell, which is probably some form of iron compound. The sulphide ores contain from 1.5 to 2 per cent of arsenic, which is present as orpiment and realgar with minor amounts of arsenopyrite. The orpiment and realgar are of a later mineralization than the gold and are barren. Roasting and washing of the sulphide ore prior to cyanidation is necessary to eliminate , the arsenic, to alter the gold coating from an insoluble form to a soluble form, and to release the gold associated with the fine sulphides. Washing the ground calcine in weak sulphuric acid, trona, or hot water are beneficial in the order named, but to date the water is the best from an economic standpoint. Metallurgical tests as well as plant practice indicate that a certain amount of gold cannot be recovered even by going beyond economic practices. On this basis metallurgical results are viewed in the light of an optimum tailing rather than a percentage of extraction. Milling Practice Crushing (Fig. I).—There are three primary ore bins, one each for oxide ore, gougy sulphide ore, and hard sulphide ore. In front of each bin a substantial yard is maintained, which makes it possible to Stockpile 2500 tons of oxidized ore, 1000 tons of hard sulphide ore, and 1000 tons of gougy sulphide ore. The bins are fed by a 60-hp. caterpillar bulldozer from the stockpiles. Fig. I shows the method and

Jan 1, 1943

By Fred Wise, C. W. Wark

The Getchell mine, a comparatively recent gold discovery, is in the old Potosi mining district, Humboldt County, Nevada. All ore is mined from open pits using Diesel shovels and gasoline trucks. Two types of ore are mined, oxide and sulphide. Oxide ores are siliceous or shaly; sulphide ores either gougy, siliceous or shaly. In mining, the sulphide ore is segregated into hard and gougy ore, because the gouge must be dried before it can be crushed. The oxide ore is not segregated. General Metallurgy The oxide ore is the leached outcrop that overlies the sulphide ore body. All of the iron and arsenic sulphides found in the sulphide ores have been thoroughly oxidized. The sulphur and arsenic have been leached out while the iron is present as hematite and limonite. The oxide ore is amenable to direct cyanidation, and although no visible free gold is present it is released at a comparatively coarse grind. The sulphide ore is more refractory, since the gold is all of micron size and some of it is locked in fine sulphides. It has been established microscopically that the free gold is coated with an impervious shell, which is probably some form of iron compound. The sulphide ores contain from 1.5 to 2 per cent of arsenic, which is present as orpiment and realgar with minor amounts of arsenopyrite. The orpiment and realgar are of a later mineralization than the gold and are barren. Roasting and washing of the sulphide ore prior to cyanidation is necessary to eliminate , the arsenic, to alter the gold coating from an insoluble form to a soluble form, and to release the gold associated with the fine sulphides. Washing the ground calcine in weak sulphuric acid, trona, or hot water are beneficial in the order named, but to date the water is the best from an economic standpoint. Metallurgical tests as well as plant practice indicate that a certain amount of gold cannot be recovered even by going beyond economic practices. On this basis metallurgical results are viewed in the light of an optimum tailing rather than a percentage of extraction. Milling Practice Crushing (Fig. I).—There are three primary ore bins, one each for oxide ore, gougy sulphide ore, and hard sulphide ore. In front of each bin a substantial yard is maintained, which makes it possible to Stockpile 2500 tons of oxidized ore, 1000 tons of hard sulphide ore, and 1000 tons of gougy sulphide ore. The bins are fed by a 60-hp. caterpillar bulldozer from the stockpiles. Fig. I shows the method and

Jan 1, 1943

By Frank C. Appleyard

Gypsum, the dihydrate form of calcium sulfate, has a history of usefulness to man dating back several thousand years, and a worldwide industry has been built on the mining and processing of this versatile industrial mineral. However, anhydrite, the other principal calcium sulfate mineral has very limited use, so much so that separate statistics for it are not always kept, but are usually included with those for gypsum. Although often found in close association, the two minerals have important chemical differences in that anhydrite is the anhydrous form of calcium sulfate (CaSO,) whereas gypsum is the dihydrate (CAS0;2H20). These two molecules of combined water result in significant differences in hardness, density, and solubility of the two minerals. However, of most importance is the difference in thermal properties wherein gypsum, upon the application of a modest amount of heat, can be converted to the hemihydrate of calcium sulfate (CaS0;Vi H,O) or plaster of paris, an intermediate product upon which is based over 90% of the value of all calcium sulfate end products. Anhydrite, on the other hand, is inert at these temperatures and its uses have been limited to only those few cases where calcium sulfate in the anhydrous form may have an advantage. The earliest known use of gypsum plaster dates back some 5000 years to the time of the Pharaohs of Egypt who had discovered that gypsum rock, when exposed to fire, broke down into a powder which when mixed with water would form a putty that could be "plastered" on a rough mud brick or stone wall to make a smooth finish. It was used down through the centuries in a limited way for ornamental purposes, but did not achieve wide acceptance because of its quick (25 to 30 min) setting or hardening time which made it difficult to use. The first real understanding of gypsum chemistry was developed in France about 1755, and the gradual growth of its present utilization dates from that time. In the United States, the first utilization of gypsum was to grind it to a fine material for use as a soil conditioner at about the time of the Revolutionary War, and the term "land plaster" for ground gypsum has remained in use. Another cementituous material-lime or calcium oxide-had become the usual plastering material, but by about 1870 a method for retarding the set of gypsum plaster was discovered and it has today almost completely replaced lime as a plastering material. In recent years, gypsum products sold in the United States have reached a total valuation of approximately $500 million per year as illustrated in [Table 1].

Jan 1, 1975

GENERAL GEOLOGIC DESCRIPTION The mineral deposit of this case study can be described as a "typical" porphyry copper deposit of the southwestern USA and northern Mexico mineral province. The copper mineralization is related to Laramide age intrusives of intermediate composition which have been intruded into older, mostly crystalline, host rocks. Post Laramide deformation and faulting have subjected the deposit to erosion, leaching, and enrichment processes, which have created a secondarily enriched chalcocite zone. Later basin and range faulting down dropped the ore zone, resulting in deep burial beneath late Tertiary conglomerates and gravels, thus preserving the deposit from further erosion and oxidation. Lower grade chalcopyrite mineralization constitutes the primary zone underlying the enriched blanket. MANAGEMENT OF SAMPLE DATA Sample data were obtained from exploration diamond drill holes that were collared vertically and drilled with surface rigs. Most of the holes were surveyed downhole and found to deviate significantly from the vertical. The data were keypunched from original drill hole logs to formats as described in Chapter 6. Three types of cards were prepared: 1) A heading card containing the following information: [Project Name Coordi-in_ Hole Collarnates Bear-clina-Total ID Northing Easting Elevinglion Depth] 2) Individual record cards following each heading card with the following information: [SampleMin. Zone interval Assay Fields Leached (enriched ID From-ToTotal Cu Oxide Cu primary)] 3) Where downhole survey data were available, a card was punched for each downhole survey point. The card image looked very much like the heading card, followed immediately after the heading card, and contained the following information: [Hole Sur Point Sur Point Sur Point ID Bearing Inclination Depth] The keypunched data were then processed by the drill hole survey program producing a record file which contained information as follows: [Line 1. Project Name Date Assay Field Identifiers Line 2. Hole ID Northing Easting Elevation Bearing Inclination Number of Samples Line 3 through N+ 2 (where N equals the number of samples in the hole) Top Min. of Min. Zone Zone Sample Sample Coordinates Sample Assays Leach InterLength Northing Fasting Elev Total Cu Oxide Cu (etc) face] MINERAL BOUNDARY PROBLEM AND TREATMENT According to the geologic description, at least three mineral zones are present within the ore limits, as well as fault boundaries and a buried topographic surface, each of which have to be considered. The interfaces of the various mineralogical and geological boundaries were identified as accurately as possible in each drill hole by the project geologist, the location of the interface was surveyed in the computer program in association with the nearest sample intercept to the interface (note last field of line 3). Thus, the boundary is represented as a point in 3D space at each intersecting drill hole. The linear interpolation (LINTERP) program then searched the last field of the surveyed record to find the top of each mineral zone, or geologic boundary, and read the coordinates associated with that sample point. The program then used the coordinates to compute an interpolated surface between all existing data points as described in Chapter 11. A contour map was plotted for each of the computed surfaces and reviewed by the project geologist. Where the surfaces were poorly defined due to lack of real drill hole information, the geologist, at times, inserted a "pseudo" drill hole intercept according to his best judgment and feeling for the configuration of the surface, and such points were entered into the data stream for LINTERP. When the project geologists were satisfied that they had obtained

Jan 1, 1980

By G. Borelius

ABOUT the turn of the century, Gibbs' thermo-dynamic theory of heterogeneous equilibrium, on the one hand, and the experimental methods of thermal and microscopic analysis, on the other, gave to the physical metallurgist his first scientific tool, the equilibrium diagram. The classical equilibrium diagram of a binary alloy system shows the boundaries between ranges of homogeneous and heterogeneous equilibrium in their dependence of concentration and temperature. A homogeneous solid sohtion which on cooling passes such a boundary is assumed to precipitate, forming a mixture of two phases with different concentrations. The equilibriunl diagram and the equilibrium theory, however, give no information about the time scheme of the process or the intermediate states passed during precipitation. For this reason it satisfies neither the practical need of the metallurgist nor the curiosity of the physicist. As a matter of fact, in the heat treatment of alloys for technical use the objective very seldom is the equilibrium state. Thus good mechanical properties of construction material are connected, for the most part, with some intermediate state. As these intermediate states are thermodynamically unstable, there is, from a theoretical point of view, always to be expected a decay of the good properties with time; and it is a matter also of practical interest to know whether this natural life time of a material is of the order of, say, ten or thousands of years. Thus, for many reasons, there is a current demand to complete our knowledge of equilibrium through knowledge of the kinetics of the precipitation phenomena. From the point of view of the physicist, the most interesting question in this case is whether there are any general laws governing the kinetics. According to a generally accepted view, precipitation is ruled by two more or less independent phenomena, the formation of nuclei of a new phase and the growth of these nuclei. It is also commonly accepted that there is a tendency for the velocity of growth to increase with increasing temperature because of the increasing mobility of the atoms. There is also a tendency for the velocity of growth to decrease in the neighborhood of the two-phase boundaries. So far, however, very little is known quantitatively about this fundamental phenomenon in the case of solid metallic systems. In our laboratory attention has been directed especially toward the nucleation phenomena, and a series of measurements have been carried out with the guidance of a work- ing hypothesis (based on experiences from previous work on order-disorder transformations in alloys) about the influence on the nucleation of thermo-dynamic potential barriers. However, before discussing the experiments, the theoretical ideas will be considered. In a binary solid solution the arrangement of atoms on the lattice points approaches with increasing temperature a state of full randomness, as illustrated by the ball model of Fig. 1, that might represent a [111] plane of a face-centered alloy with 30 pct "black" and 70 pct "white" atoms. In reality the atoms are changing places continually with their neighbors so that the picture should rightly have been a moving one. On account of this thermal motion the concentration of black atoms within a certain group of, say, a hundred or a thousand lattice points fluctuates with time around the bulk concentration of 30 pct in a manner governed by statistical laws. With decreasing temperature two independent changes in this state grow more and more important. First, the mobility of the atoms decreases, and second, the forces between the atoms will have an increased influence on the fluctuations. In alloys with a tendency for precipitation, which are the concern of this lecture, the distribution function of concentration fluctuations will broaden, so that the relative probability of great local variations from the bulk concentration increases. Fig. 2 gives an example of such a fluctuation. When the alloy is supercooled below the solubility limit into the range of two-phase equilibrium, the fluctuations will now and then at some point give rise to a state that resembles the equilibrium state and thus will form a stable nucleus that is capable of growing by diffusion processes. In discussions with colleagues and in the literature, I have often encountered the idea that three or four atoms of the dissolved metal could form a nucleus of the new phase. A look at the ball model might be enough to indicate that this cannot be true. If it were true, there should be nothing but nuclei, whereas we know from experiment that nucleation must be a rather rare occurrence. In fact we have, as will be mentioned later, certain reasons to believe that the nuclei are formed by fluctuations containing some hundreds of atoms, which should be the order of the number of black balls in the fluctuating group of the figure, if it were extended into three dimensions. As a working hypothesis we have assumed that the fluctuations producing nuclei, though large and rare, still are ruled by the distribution laws of fluctuations of the supercooled solid solution in its initial state. Thus the probability of nucleation will be connected to the thermodynamic properties of the solid

Jan 1, 1952

By P. Ramachandraraa, T. R. Anantharaman

THE technique developed by Duwez, Willens and Kle-ment' for rapid solidification of molten alloys in small quantities by ejecting them on to a highly conducting substrate with the aid of a shock wave has come to be known as "liquid quenching" or "splat cooling" in literature and has already been applied to alloys of many binary systems.' The very high cooling rates of the order of l060 to 1080 per sec obtained in this technique and the consequent drastic undercooling of the melt have led to such striking results as formation of new amorphous and crystalline metastable phases and extension of solid-solubility limits in many alloy systems. The cooling rates in these experiments can, however, be neither accurately measured nor effectively controlled and hence the occurrence of either more than one metastable phase or appreciable differences in constitution in splat-cooled alloys of the same composition cannot be ruled out.3 In fact, workers in two laboratories have reported formation of two different cubic metastable phases in liquid-quenched Au-Si alloys containing 17 to 20 at. pct si.4,5 The present note deals with some new and interesting results obtained in a Ag-21 at. pct Ge alloy and a Au-27 at. pct Ge alloy subjected to the liquid-quenching treatment at different cooling rates. Two alloys of nearly the same composition had earlier been rapidly solidified by the standard "gun technique" in the same laboratory and shown to solidify in nonequilibrium structures, viz., an hcp phase (a = 2.987Å;c =4.718Å) in the Ag-Ge system6,7 and both hcp (a = 2.885Å; c = 4.754Å) and tetragonal (a = 11.627A; c = 22.491Å) phases in the Au-Ge system.3 Apart from different substrates (copper and stainless steel), different melt temperatures (50º to 500°C above the liquidus), and different capillary sizes (0.4 to 1.5 mm diam) in the "gun technique", slower cooling rates of the order of l050C per sec were also obtained in the present work by replacing the substrate by water at room temperature in a modification of the technique due to Olsen and Hultgren.8 No attempt was made to measure the cooling rate in any experiment. The products varied from fine powder (>40 pct by wt—200 mesh) obtained by quenching in water to thin flakes and foils (11.0 sq mm to 4.0 sq cm) produced by splat cooling. They were subjected to the Debye-Scherrer method of X-ray examination in each case. In case of the Ag-Ge alloy, in addition to the expected metastable hcp phase,6 a coexistent tetragonal phase (a = 5.093,Å; c = 7.009Å) was detected in some experiments. Eight weak X-ray reflections from this phase could be identified, two of them being the same as reported earlier.7 With increase in the cooling rate, the dominant hcp phase seems to be first accompanied by traces of silver or germanium and then either homogeneous or mixed with traces of this new tetragonal phase. The splat-cooled Au-Ge alloy showed generally the expected hcp and tetragonal phases,3 but sometimes traces of germanium were also present. Further, two new tetragonal phases with the following lattice param ters could be clearly identified in samples quenched

Jan 1, 1970

By N. S. Greensfelder

THE firing of explosive charges by electricity dates back to 1745 when a Doctor Watson is said to have used an electric spark for igniting gunpowder. His method failed in practical application because, to ignite blasting powder, heat must be applied for a period longer than the duration of an electrostatic spark; Wheatstone found that this does not exceed 14000 sec. The usual result, therefore, of an electrical discharge in gunpowder is to scatter the grains, without igniting them. Benjamin Franklin partly overcame this difficulty by confining the powder and crushing grains that lay in the line of discharge. In 1822, after the primary cell had been discovered by Volta, Robert Hare, professor of chemistry at the University of Philadelphia, noticed the heating of a short piece of wire placed in the circuit of a voltaic battery. Later work by Oersted, Faraday, Wheatstone, and Henley resulted in the development of the magneto blasting machines. Fulminates of gold and silver are thought to have been discovered by Cornelius Drebbel, about 1600. Pepys, in his diary for November 11, 1663, recounts a conversation with Doctor Allen, who told him about aurium fulminans "of which a grain ... . put in a silver spoon and fired, will give a blow like a musquett and strike a hole through the silver spoon downward, without the least force upward.1," This fallacious belief that explosives act downwards only, is still held by many who should know better. Fulminate of mercury was discovered by Howard in 1799. J. Shaw of Philadelphia is said to have invented a fulminate-of-mercury cap in 1814; his idea was later adopted by gunmakers in England and France. The principal advantages of

Jan 9, 1922

By Gilbert E. Doan

IN a technical civilization, that is. one whose major difference from past civilizations is its enormous development of technology, in transportation, communication, labor saving, centralized control, and mass production, it is natural that the most critical problems should be those that arise out of technologic advances; it is natural also that a. man who understands these advances should be chosen as the leader to meet these problems. This was done in the selection of Herbert Hoover as President of the United States. Mr. Hoover was a good, and perhaps a great President. But he was not a great politician and that ensured his defeat when a crisis came.

Jan 1, 1939

By Charles T. Holland

Miners have observed for many years that as pillars are removed in mining operations the pillar adjacent to those mined out frequently shows evidence of being compressed. Although this has been known for a long time, little or no information is available concerning how far from the mining area the pillars are affected in this manner, or how great is the magnitude of the compression, or concerning the manner or characteristic way in which the compression acts, as the mined-out area approaches the pillar under consideration. The writer made a study of pillar action in the Mona mine of the Arkwright Coal Co., partly to get some quantitative data on the deformation of coal-mine pillars. The study was not completed but it seems best to present the information that was obtained. Method of Measurement Fig. 1 shows the details of the apparatus used for measuring the deformation. This is permanently attached to the pillar. Fig. 2 shows the details of the gauge that records the deformation. This gauge is removable and was used to measure the deformation at all the stations. The recording dial is graduated in thousandths of an inch and records measurements up to one inch. Fig. 3 shows how the removable gauge is placed in the instrument. When installing the instrument in the side of the pillar, all loose coal near by was removed and a trench 4 to 6 in. deep, and approximately 1 ft. wide, was dug in the pillar. In this trench near the top and the floor, the holes holding the instrument were bored. These holes were then partly filled with thin cement and the rods were inserted. The remaining space in the holes was tamped full of cement. At this time, by carefully manipulating the permanent rods and clamps and adjustable pin, it was possible to adjust the instrument so that the removable gauge would stand true and measure any vertical deformation between the top and bottom rods. Fig. 4 is a photograph of one of the instruments installed, showing the removable gauge in place.

Jan 1, 1938

By Charles T. Holland

Miners have observed for many years that as pillars are removed in mining operations the pillar adjacent to those mined out frequently shows evidence of being compressed. Although this has been known for a long time, little or no information is available concerning how far from the mining area the pillars are affected in this manner, or how great is the magnitude of the compression, or concerning the manner or characteristic way in which the compression acts, as the mined-out area approaches the pillar under consideration. The writer made a study of pillar action in the Mona mine of the Arkwright Coal Co., partly to get some quantitative data on the deformation of coal-mine pillars. The study was not completed but it seems best to present the information that was obtained. Method of Measurement Fig. 1 shows the details of the apparatus used for measuring the deformation. This is permanently attached to the pillar. Fig. 2 shows the details of the gauge that records the deformation. This gauge is removable and was used to measure the deformation at all the stations. The recording dial is graduated in thousandths of an inch and records measurements up to one inch. Fig. 3 shows how the removable gauge is placed in the instrument. When installing the instrument in the side of the pillar, all loose coal near by was removed and a trench 4 to 6 in. deep, and approximately 1 ft. wide, was dug in the pillar. In this trench near the top and the floor, the holes holding the instrument were bored. These holes were then partly filled with thin cement and the rods were inserted. The remaining space in the holes was tamped full of cement. At this time, by carefully manipulating the permanent rods and clamps and adjustable pin, it was possible to adjust the instrument so that the removable gauge would stand true and measure any vertical deformation between the top and bottom rods. Fig. 4 is a photograph of one of the instruments installed, showing the removable gauge in place.

Jan 1, 1938

By John C. Ludlum

A most stimulating event in the development of world mineral production and technology, the Eighth Commonwealth Mining and Metallurgical Congress was convened at Melbourne Town Hall on March 1 by Sir Maurice Mawby, President of the Congress. Continuing successively at Melbourne, Canberra, Sydney, and Surfers Paradise in Australia and then at Wellington, New Zealand, the four plenary sessions, 50 technical sessions, and both local and extended mine and plant tours during the seven weeks of the Congress, provided a superb medium for assessing the present and looking into the future. There were more than 2300 participants from 32 countries in the various functions of the Congress, including 500 delegates and wives from countries other than Australia and New Zealand, and from 18 countries not members of the British Commonwealth of Nations. Registrants from South Africa numbered close to 200 and there were about 100 from Canada, 70 from the United Kingdom, 60 from the United States, and 20 from Zambia. Although most of the overseas delegates held administrative positions with companies, government agencies, and colleges or universities, professions identifiable from registrations included 100 engineers, 60 geologists and 30 metallurgists.

Jan 8, 1965