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By A. P. Pipilen, M. Weintraub, W. F. Hosford, A. A. 230-000-000-006 Orning

In a study of the principal factors affecting the transport of coal-water mixtures through a centrifugal pump and a pipeline, the interrelation between solids concentration, velocity, and pressure drop was established for a minus 2-in. bituminous lump coal in concentrations up to 48% by weight. Optimum concentrations were found to exist for maximum capacity of a given pipeline and for minimum energy requirements per ton of coal. Coal size degradation by particle fracture took place in the pump; abrasion and attrition to form fines took place in both pump and line. A novel system of studying flow of solid-liquid mixtures in pipelines by the determination of relative sound intensities was developed. The hydraulic transport of minerals is in part an established art and in part a relatively new field of industrial technology. The established art developed especially where the mineral was under water, where water was used in hydraulic mining, or where large quantities of water had to be pumped from the mining site to expose the mineral. Power requirements and limited life of pumps and pipelines, under such conditions, were relatively unimportant. Examples may be found in dredging, and in hydraulic mining and transportation of phosphate rock.' In the design of a coal pipeline the most needed information relates to flow characteristics, power requirements, pressure drop, size degradation, and size segregation as a function of solids-water ratio and size distribution of the mineral. A considerable amount of information has become available concerning pressure drop and energy requirements.2'3 These topics were therefore not studied intensively. Coal degradation, however, has not been clearly delineated. Lammers and coworkers4 showed that coal may be severely degraded in size but did not distinguish between degradation produced in the pump and in the line. It is the primary purpose of this paper to outline techniques used to acquire some of the much needed design data. A full report of the investigation is to be published by the Bureau of Mines at a later date. APPARATUS AND PROCEDURE The apparatus for circulation tests (hereafter referred to as Type III tests) consisted essentially of a loop and a centrifugal solids handling pump, with attendant instrumentation, and coal weighing and feeding equipment. It also included an adequate water supply, as shown in Fig. 1. The selection of 6-in., schedule 40, carbon steel pipe was based on the top size of coal to be used. Assuming that the top size of coal should not exceed 1/3 the inside diameter of the pipe, a 6-in. pipe would accommodate 2-in. coal, a realistic size for transport from the mine face. An 872-ft loop was constructed first. Later a few tests were made with a lengthened circuit of 1361 ft by superimposing an additional loop on the original one. All bends used were of 3-ft radius. A noisemeter microphone, taped to the pipe at various points, but generally about 50 ft downstream of the pump, was used to sense segregation of the coal and the length of slugs of coal in water. To measure the stream velocity, radioactive sources were encapsulated in plastic cubes. The specific gravity of the blocks was adjusted to 1.3 and 1.8 to simulate low ash and high ash coal particles. Their passage through the pipeline was detected by a Geiger counter. Because the repeated circuits of the Type III tests did not permit coal degradation in the line to be distinguished from that produced by the pump, the two

Jan 1, 1964

By Herbert Kilburn Scott

The information in this paper was collected by the author during a five years' residence in the State of Minas Geraes. Outside the State itself, very little is known of the gold-field. Such accounts as have been published are antiquated and contained in publications mostly out of print. The increased attention now attracted to the natural resources of Brazil, and the probability that American capital will be, in the near future, largely invested there, will make an up-to-date account specially useful at this time. I. Historical and General. Soon after the Jesuits founded what is to-day the city of S?o Paulo, gold was discovered in the streams of the neighborhood; and this incited the natives, known as paulistas, to look for it farther afield. After many expeditions into the interior, gold was found in quantity in placer-deposits in the vicinity of Ouro Preto, in 1699, by a prospecting party headed by Antonio Dias. The name of Ouro Preto (black gold) was given because of the superficial dark coating of the gold, and the inhabitants of the district were known as mineiros (miners). Later on, the placers of Cuiaba, in the State of Matto Grosso, were discovered in 1718, and those of the State of Goyaz in 1724. All this territory was at the time a private colony, but on account of its increasing importance it was taken over by the Crown, and in 1739 a royal governor was appointed, with his nominal capital in S?o Paulo, though for a number of years the real capital was Ouro Preto. In 1711 Ouro Preto was raised to the rank of a city, with the name of Villa Rica de Ouro Preto (" rich city of black gold"); and in 1720, by reason of administrative difficulties, Minas Geraes was constituted a separate colony from S?o Paulo, with Villa Rica as its capital. The site of this city

Jan 1, 1903

By R. A. Burmeister, P. E. Greene

Two methods for the synthesis and growth of single crystals of B6 P are described: a solution-growth process employing nickel as the solvent, and a vapor-growth process employing BBr3 and pH3 as the reac-tants. Relevant portions of the Ni-B-P phase diagram have been examined and liquidus data are presented for an isobaric section at an equivalent P2 pressure of 3.8 Torr, and an isothertnal section at 1300°C. Solution-growth techniques are described and the purity and crystalline perfection of the resulting material is discussed. Epitaxial growth of BeP on &P substrates has been obtained using the vapor-phase system. Evidence is presented which supports the assignment of the stoichioknetry B6P to boron sub-plwsphide. THE B-P system differs from other III-V systems in that it contains two distinct compounds. The existence of the monophosphide (BP) has been known for some time,' whereas the subphosphide (B6P) was reported only recently.' Both of these compounds are large bandgap semiconductors (Eg = 2.0 ev for BP~ and 3.3 ev for B6P4), and are therefore of interest for optoelectronic applications. The crystal structure of B6P is quite different from those of most semiconductors, being derived from that of a-rhombohedral boron by the incorporation of the two phosphorus atoms in the intericosahedral sites of the a-boron unit cell?76 B6P also exhibits a number of unusual characteristics, including a high degree of chemical inertness, hardness, and high-temperature stability. No information concerning other physical properties is available in the literature. The present study is part of a broader investigation of the preparation and properties of B6P. This paper considers methods of growing single crystals of B6P of suitable size and quality for physical-property measurements. A detailed discussion of the electrical and optical properties of B6P will be given elsewhere.' GENERAL CONSIDERATIONS The method first used to synthesize B6P was the disproportionation of BP' according to the reaction The equilibrium P2 pressures corresponding to this reaction were also determined .'" It is therefore possible, in principle, to apply several of the reactions reported for the synthesis of BP to the synthesis of B6P merely by selecting lower phosphorus pressures or higher temperatures, in accordance with the equilibrium data. Other techniques for the direct synthesis of microcry stalline B8P have also been reported.1°

Jan 1, 1968

By Andrew Mayer

Early in 1942 National Lead Co. was requested by the War Production Board to construct and operate a plant for the Government to produce magnesium by the ferrosilicon process which had been developed by Dr. E. M. Pidgeon in the laboratories of the Canadian National Research Council. A contract with the Defense PlantCorporation was concluded in May 1942 and Magnesium Reduction Co., a wholly owned subsidiary of National Lead Co., was formed to carry out this project. The capacity of the plant was rated at I0,000,000 lb. of magnesium per year, equivalent to an average production of about 14 tons per day. The Process The process, in brief, is as follows: The raw materials are dolomite and ferro-silicon, 75 per cent grade. The dolomite is calcined, the calcine and ferrosilicon are ground and mixed and the mixture is briquetted. The briquets are charged into tubular retorts of chrome-nickel steel set horizontally in a furnace with the open ends projecting outside the front wall. The retorts are then closed and evacuated. Magnesium is liberated according to then reaction 2(MgOl CaO) + Si = 2Mg + zCaO, Si% It is distilled from the charge and con- densed on a removable sleeve in the throat of the retort. Sodium and potassium, if present, are also liberated, distilled and condensed on a "sodium condenser" in the end of the retort. At the expiration of the distillation period the retorts are discharged and the cycle of operations is repeated. For the rated daily capacity of 14 tons of magnesium there are required approximately 170 tons of dolomite, equivalent to about 85 tons of calcine, and 16.5 tons of ferrosilicon. The retort residue, a bulky powder which at present is waste, amounts to about 87 tons. Choosing the Site The investigation to select the location of the plant was conducted by the Research Laboratories of National Lead Co. This work included geologic and economic surveys of a number of districts, examination of dolomite samples by chemical, spectrographic and petrographic methods and large-scale tests of dolomite from several of the more promising localities. in the last-named tests 2-ton samples were put through the ferrosilicon process in the pilot plant of the Canadian National Research Council, under the guidance of Dr. Pidgeon. The conclusions drawn from the investigation, in regard to the quality of dolomite to be used in the ferrosilicon process, were: First, the dolomite should contain not less than 21 per cent MgO and not more than 2 per cent acid-insoluble material. Second, the dolomite should

Jan 1, 1944

By M. E. Wadsworth, J. S. Cho

The kinetics of silicate adsorption from aqueous colloidal (SO2 . nH2O) silicate sols on thoria (ThO 2) surfaces were determined at 100, 150 and 200°C in an autoclave system. The amount of chemisorbed silica on the thoria surface was determined by means of infrared spectroscopy. First order kinetics were observed in all cases with the anomalous result that the first order rate constant was found to vary with the initial amount of silica added to the system. These results have been correlated quantitatively by means of a strain model of adsorption involving two types of surface silicate bonding, both of which may be readily observed in the infrared spectrum. Initial adsorption occurs rapidly to produce one type of surface silicate followed by a slow transition to a second type of surface bonding, the latter transition representing the slow step in the adsorption process. Initial interest in silicate adsorption on thoria arose from the dispersive effect silica has on finely divided thoria in aqueous suspensions. Spaepen et al studied the equilibrium adsorption of silica on thoria and noted increased adsorption up to approximately 215°C with a notable decrease in adsorption at temperatures above 215°C. Similar results were observed by Cho and Wadsworth2 who identified two types of surface silica by infrared spectroscopy. At temperatures of 100°C and above, rapid initial adsorption occurred producing a surface species with maximum infrared absorbance at 9.4 After extended periods of time, the 9.4 species converted to a more strongly bonded surface silicate with maximum infrared absorbance at 10.3. The latter was very strongly bonded and difficult to remove by desorption in hot distilled water, while the 9.4 species could be removed more readily by similar treatment. The 9.4 and 10.3-p species were distinguishable from free unadsorbed silica which had maximum infrared absorbance at 9.1 The kinetics of the formation of the strongly chemisorbed silica (10.3) on thoria as determined by infrared examination are presented in this study. EXPERIMENTAL PROCEDURE As a source of silicate, Ludox SM (16.36% SiO2) or Ludox HS (30.25% SO2), DU Pont colloidal silica in aqueous suspensions, was added directly to a thoria slurry made up of one liter of distilled water and one gram of thoria. The thoria used was designated D-16-650 and was received from the Chemical Technology Div. of Union Carbide Nuclear Co. It was prepared by firing thorium oxalate at 650°C. The D-16 thoria had a surface area of 32.4 ± 0.6 sq m per g as determined by nitrogen adsorption.* This thoria 'Determined by Unlon Carbide Nuclear Co., ORNL. had a total adsorption capacity for silicate of approximately 19.4 mg of silica per gram of thoria.1 This amounts to approximately one SiO2 molecule for every surface thoria site. In all cases in this study.

Jan 1, 1963

By C. Pin, J. Bruce Wagner

In order to demonstrate the role of defect chemistry in the solid state to a process-metallurgy reaction, laboratory experiments were carried aut to remove copper from liquid lead using lead sulfide which had been doped to provide definite concentrations of cation vacancies. The removal of copper decreases in the order: bismuth- or antimony-doped PbS, undoped PbS or tin-doped PbS, and silver-doped PbS. The data are explained using the simple mass -action law and in terms of the exceedingly rapid interstitial diffusion of coppw in PbS to vacant cation sites. Results are in agree-ment with thermodynamic predictions and certain industrial practices that copper may be removed from lead below the limit of its coexistence with pure Cu,S and pure PbS. The role of defects in solids has been exploited in semiconductors and many device applications, in explaining tarnish formation, dissolution rates, and mechanical properties. The application of these principles to many process-metallurgical problems has not received as much attention. It was the purpose of this study to illustrate the role of defect structures of solids in a process-metallurgical application. When lead is refined, the ore, galena (PbS), is reduced by a series of operations to liquid lead (lead bullion) containing impurities such as copper, bismuth, antimony, arsenic, and silver depending on the original composition, the fluxes added, and the particular type of operation, see Fig. 1. In order to produce a pure lead, a series of refining operations are carried out. A typical lead-bullion analysis before drossing is shown in Table I. Copper is removed first as its presence increases the zinc consumption during the desilverizing process. For preliminary copper removal the lead bullion is brought to a temperature of 440°C in a kettle of 60 metric tons capacity. Some tar (or sawdust or resins) is usually added and the kettle is cooled slowly to 400°C while the liquid is stirred. A solid phase, called a dross, floats on top of the liquid and the particles of the solid phase adhere to the tar. This "hot dross" is collected. It contains mostly lead and copper which are partly oxidized and/or sulfurized if the lead bullion contains some sulfur.*

Jan 1, 1963

By Carl J. Christensen

This paper presents the development of diamond coring of the Weber sand section in the Rangely Field, Colorado. The description and operation of the diamond-coring equipment is included as well as the economics and results obtained by its use. Diamond coring is compared with the other methods used for drilling the Weber sand. Two types of diamond coring are discussed. The first and the one used most extensively is the regular 6 ?x 3¾-in. diamond bit and bottom hole 50-ft barrel. The second type, which has also been used successfully by Stanolind, consists of a 4 13/16 x 2 11/32-in. cutting bit and reverse circulation coring equipment. The first type will be referred to as regular or conventional diamond coring while the latter type will be referred to as reverse circulation diamond coring. Introduction Diamond coring has been used for many years in mines, quarries, surface coring at dam sites, laboratory work, and so on, but, until recently, had never been used very extensively in coring oil sands and other oil-bearing strata. It was demonstrated early in its development that hard forma-tions could be cut rapidly and that core recoveries far exceeded those obtained with conventional hard rock bits and quite often 100 pet recovery was obtained. However, the poorly constructed diamond-coring equipment and resulting high costs of early diamond coring slowed down its progress. Early in the development Of the Rangely Weber field, Rio Blanco County, Colorado, the Rangely Engineering Committee selected key wells that were to core the Weber section in its entirety. These cores were to be analyzed and various sections tested to gather sufficient information so that accurate reservoir studies could be made. The number of wells selected were kept at a minimum inasmuch as the productive Weber section, which is found at an approximate depth of 6000 ft, is thick, being as much as 600 to 700 ft on top of the structure. The Weber section is also comparatively hard to drill. For instance, one well took 60 days to core 490 ft of Weber sand with conventional coring equipment. It was noticed, however, that Core Labora-verse diamond plug cutter could quickly cut a plug from. one of these hard cores. This fact led to the serious consideration of diamond coring the Weber sand. The first diamond coring at Randy was performed by Stanolind in August 1946 andl although we experienced some difficulties, the results definitely indicated that diamond coring would be successful at Rangely for .coring the Weber sand. As more wells were cored and the equipment improved somewhat, diamond coring became less expensive than drilling the Weber sand with the conventional rock bit. In less than one Year diamond coring has definitely Proved its usefulness in oil-field drilling operations in the Rocky Mountain area where formations quite often drill very hard. The practice of spot coring in hard oil-bearing formations with conventional coring equipment is giving way diaimond coring of the entire sections, thus givng an operator the complete picture.

Jan 1, 1948

By A. D. Hughes

A relatively small area in Malaya, about 200 miles long by 40 miles wide, is the most important source of tin in the world. Some tin is recovered in other parts of the peninsula. Of the tin mined, 98 pct is recovered from alluvial deposits. From 1935 to 1941 the average annual world production of tin was 190,000 tons. The average annual production from Malaya during the same period was 62,000 tons or about one third of the total. Other producing countries, in order of importance, were the Netherlands East Indies 34,000 tons, Bolivia 30,000, Belgian Congo, Nigeria, Siam, Burma, China, and a few others with smaller amounts. The serious shortage of tin during the war period was due to the fact that the Japanese were occupying Malaya, Netherlands East Indies, Siam, Burma, and China which, together, formerly were producing 65 pct of the world supply. Because of the importance of tin in the world economy, and because of the fact that little is generally known in this country regarding the methods of recovering the tin in Malaya, it is believed that a short description of the industry, as carried on there, may be of interest to many of the readers. In addition you may be interested in the effect of the Japanese occupation, subsequent recovery of the properties and conditions existing at that time. The Malay Peninsula lies at the southeast tip of Asia. Singapore, a British colony, is situated on a small island at the southern end of the peninsula. It is a city of nearly a million population, an important world port, the site of a large British naval base and of one of the tin smelters. The Federation of Malaya consists of nine states, plus Penang, Malacca, and other small areas formerly part of the Straits Settlements, and it now covers the entire peninsula. The civil administration is controlled by the British. A constitution was adopted recently after reaching agreement with the Malay Sultans of the various states, who retain control of religious matters. The Capital is at Kuala Lumpur, a city of about 200,000 population, 250 miles north of Singapore and about half way to Penang. The country covers an area of 51,000 square miles, an area about equal to that of Florida and about one third that of California. It has a population of about seven million consisting of Malays, Chinese, East Indians, and a relatively small number of Europeans. The principal products are tin and rubber. There are over three million acres of planted rubber. The country is not self-supporting agriculturally as about one third of the necessary food supplies is imported. Rice is the principal food of the Asiatic population and some of the recent unrest is caused by a shortage of rice and the consequent high price. Good highways have been built through the settled part of the country. A railway, Government owned, extends from Singapore northward through the entire peninsula. The climate is tropical but health conditions are good for a tropical country. I was in Malaya from 1939 until January 1942, during the time of the simultaneous air attacks on Pearl Harbor, Manila, Hongkong, and Singa-

Jan 1, 1950

By Carl J. Christensen

This paper presents the development of diamond coring of the Weber sand section in the Rangely Field, Colorado. The description and operation of the diamond-coring equipment is included as well as the economics and results obtained by its use. Diamond coring is compared with the other methods used for drilling the Weber sand. Two types of diamond coring are discussed. The first and the one used most extensively is the regular 6 ?x 3¾-in. diamond bit and bottom hole 50-ft barrel. The second type, which has also been used successfully by Stanolind, consists of a 4 13/16 x 2 11/32-in. cutting bit and reverse circulation coring equipment. The first type will be referred to as regular or conventional diamond coring while the latter type will be referred to as reverse circulation diamond coring. Introduction Diamond coring has been used for many years in mines, quarries, surface coring at dam sites, laboratory work, and so on, but, until recently, had never been used very extensively in coring oil sands and other oil-bearing strata. It was demonstrated early in its development that hard forma-tions could be cut rapidly and that core recoveries far exceeded those obtained with conventional hard rock bits and quite often 100 pet recovery was obtained. However, the poorly constructed diamond-coring equipment and resulting high costs of early diamond coring slowed down its progress. Early in the development Of the Rangely Weber field, Rio Blanco County, Colorado, the Rangely Engineering Committee selected key wells that were to core the Weber section in its entirety. These cores were to be analyzed and various sections tested to gather sufficient information so that accurate reservoir studies could be made. The number of wells selected were kept at a minimum inasmuch as the productive Weber section, which is found at an approximate depth of 6000 ft, is thick, being as much as 600 to 700 ft on top of the structure. The Weber section is also comparatively hard to drill. For instance, one well took 60 days to core 490 ft of Weber sand with conventional coring equipment. It was noticed, however, that Core Labora-verse diamond plug cutter could quickly cut a plug from. one of these hard cores. This fact led to the serious consideration of diamond coring the Weber sand. The first diamond coring at Randy was performed by Stanolind in August 1946 andl although we experienced some difficulties, the results definitely indicated that diamond coring would be successful at Rangely for .coring the Weber sand. As more wells were cored and the equipment improved somewhat, diamond coring became less expensive than drilling the Weber sand with the conventional rock bit. In less than one Year diamond coring has definitely Proved its usefulness in oil-field drilling operations in the Rocky Mountain area where formations quite often drill very hard. The practice of spot coring in hard oil-bearing formations with conventional coring equipment is giving way diaimond coring of the entire sections, thus givng an operator the complete picture.

Jan 1, 1948

By M. R. Pickus, I. W. Pickus

Ordinarily age-hardening is thought of as being associated with a limited solubility of one metal in another. Much less has been written about the type of age-hardening that attends the formation of superstructures. Solid solutions generally have a disordered or random arrangement of atoms; that is to say, in a unit cell, such as that of the face-centered cubic lattice, there is no distinction between atomic sites. Atoms of a given metal may occupy cube corners or face centers. Under certain conditions, however, some alloys develop what is known as an ordered structure. In this case, cube corners would be occupied by only one type of atom and face centers by another type. Associated with this ordering process are many interesting changes in physical and electrical properties. Moreover, in many instances the alloys become susceptible to age-hardening. From the evidence now to be described, it appears that age-hardening in 14-carat gold alloys is due to the formation of both a superstructure and a precipitate. Alloys of the metals gold, silver and copper were known in antiquity. The available knowledge regarding their constitution has been ably summarized and extended by Sterner-Rainer,' Jänecke,2 Carter,3 Wise,4 and Coleman.5 Despite the long familiarity with these alloys and the numerous investigations of their properties that have been made, there is still much that remains unexplained. It has been known for a long time that the hardness of these alloys is closely related to the rate of cooling. Similar changes have been observed in the binary gold-copper systemI6 one alloy of which corresponding to the formula CU³AU, has been the subject of a careful and competent report by the English investigators, Sykes and Evans.7 They have established the fact that the shift of atoms from the ordered to the disordered condition takes place in some critical temperature range. At temperatures above the critical range the structure is random, and at temperatures below it the structure is ordered. No difference between these two structures is revealed by the microscope. Both have a face-centered cubic lattice and complete solubility, so that there is no evidence to suggest that the changes in properties are associated with a phase change. The more slowly the alloy is cooled, the higher is the degree to which the ordering process proceeds and the more marked is the increase in hardness. As experimental data were accumulated on I4-carat gold alloys, several contradictions arose. Certain observations tended to confirm the existence of a superstructure, while other observations seemed to be inconsistent with such a transformation. Present Investigation As a focal point about which to develop this subject, it is expedient to consider first

Jan 1, 1943

By M. R. Pickus, I. W. Pickus

Ordinarily age-hardening is thought of as being associated with a limited solubility of one metal in another. Much less has been written about the type of age-hardening that attends the formation of superstructures. Solid solutions generally have a disordered or random arrangement of atoms; that is to say, in a unit cell, such as that of the face-centered cubic lattice, there is no distinction between atomic sites. Atoms of a given metal may occupy cube corners or face centers. Under certain conditions, however, some alloys develop what is known as an ordered structure. In this case, cube corners would be occupied by only one type of atom and face centers by another type. Associated with this ordering process are many interesting changes in physical and electrical properties. Moreover, in many instances the alloys become susceptible to age-hardening. From the evidence now to be described, it appears that age-hardening in 14-carat gold alloys is due to the formation of both a superstructure and a precipitate. Alloys of the metals gold, silver and copper were known in antiquity. The available knowledge regarding their constitution has been ably summarized and extended by Sterner-Rainer,' Jänecke,2 Carter,3 Wise,4 and Coleman.5 Despite the long familiarity with these alloys and the numerous investigations of their properties that have been made, there is still much that remains unexplained. It has been known for a long time that the hardness of these alloys is closely related to the rate of cooling. Similar changes have been observed in the binary gold-copper systemI6 one alloy of which corresponding to the formula CU³AU, has been the subject of a careful and competent report by the English investigators, Sykes and Evans.7 They have established the fact that the shift of atoms from the ordered to the disordered condition takes place in some critical temperature range. At temperatures above the critical range the structure is random, and at temperatures below it the structure is ordered. No difference between these two structures is revealed by the microscope. Both have a face-centered cubic lattice and complete solubility, so that there is no evidence to suggest that the changes in properties are associated with a phase change. The more slowly the alloy is cooled, the higher is the degree to which the ordering process proceeds and the more marked is the increase in hardness. As experimental data were accumulated on I4-carat gold alloys, several contradictions arose. Certain observations tended to confirm the existence of a superstructure, while other observations seemed to be inconsistent with such a transformation. Present Investigation As a focal point about which to develop this subject, it is expedient to consider first

Jan 1, 1943

By A. J. May

THE Star shaft is near the north boundary of the group of mining claims belonging to the Ground Hog Unit of the American Smelting and Refining Co., near Vanadium, N. Mex. The shaft bins and surface plant are served by a siding from the Atchison, Topeka and Santa Fe railroad, and this factor of railroad transportation influenced the choice of site. This paper describes the shaft sinking only, and does not cover the surface plant or other parts of the mine or its equipment. A. J. MAY is Engineer in Charge of Shaft Sinking, Ground Hog Unit, American Smeling and Refining Co, Vanadium, New Mexico. El Paso Meeting, October 1948. TP 2737 A. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Nov. 29, 1948; revision received June 22, 1949. Experience with this shaft seems of interest because of the mechanical mucking, good rate of progress, and the costs obtained at the present level of labor and material costs. The shaft has four compartments, as shown in fig. 1. Equipment: Riddell Mucker: The mucking was done by a clamshell bucket and a Riddell mucker, which, as in other shafts sunk with this machine, consisted of a carriage mounted on axles with double flanged wheels that traveled back and forth on a swinging track, parallel to the long axis of the shaft (fig. 2 to 6). Three air motors were mounted on the carriage; one hoisted the clam, one operated the bucket opening and closing device, and the third moved the carriage along the track. The carriage was designed with guide shoes and could be hoisted to the surface in either skip compartment without trouble (fig. 4). During blasting operations the clam was tied on the rails and the carriage was hoisted up a few sets to protect it from flying rock. No attempt was made to hoist the carriage and clam together lest the clam catch in the timbers. The track on which the carriage operated was built of extra heavy 6 in. pipe with 1x2 in. strap welded on top for rails. A Yzx4 in. angle was welded on the bottom of the 6 in. pipe with a 1 in. round rod welded in the angle for a filler and additional strength. The end or cross pieces of the track which kept the track to gauge, were also made of reinforced, heavy 6 in. pipe and were bolted to the long members on which the carriage traveled (fig. 2 and 3). This section of track was used, instead of heavy railroad rails, for two reasons: (1) this section would bend before it snapped and was much safer for the men in the bottom; and (2) it was a better blasting shield, and protected the timbers better, in fact, only five wall plates were broken by blasting in 1926 ft of shaft, although 150 to 225 lb of 40 pct powder was loaded per round, depending on the ground conditions. The swinging track was supported by four % in. hoisting cables anchored to the end plates. There were also four safety cables fastened two to three sets above the carriage, which would catch the carriage if one of the supporting cables should fail for any reason (fig. 7). Clamshell Buckets: Two clams were used, each of % yard capacity, both manufactured by Blaw-Knox, Nos. 662 and 663. The No. 663 clam was 4 in. wider and 5 in. shorter than the No. 662, but seemed to dig and load more satisfactorily. It was easier to hoist and lower when it came up for repair. It did not turn over as easily, had less spill

Jan 1, 1951

By A. J. May

THE Star shaft is near the north boundary of the group of mining claims belonging to the Ground Hog Unit of the American Smelting and Refining Co., near Vanadium, N. Mex. The shaft bins and surface plant are served by a siding from the Atchison, Topeka and Santa Fe railroad, and this factor of railroad transportation influenced the choice of site. This paper describes the shaft sinking only, and does not cover the surface plant or other parts of the mine or its equipment. A. J. MAY is Engineer in Charge of Shaft Sinking, Ground Hog Unit, American Smeling and Refining Co, Vanadium, New Mexico. El Paso Meeting, October 1948. TP 2737 A. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Nov. 29, 1948; revision received June 22, 1949. Experience with this shaft seems of interest because of the mechanical mucking, good rate of progress, and the costs obtained at the present level of labor and material costs. The shaft has four compartments, as shown in fig. 1. Equipment: Riddell Mucker: The mucking was done by a clamshell bucket and a Riddell mucker, which, as in other shafts sunk with this machine, consisted of a carriage mounted on axles with double flanged wheels that traveled back and forth on a swinging track, parallel to the long axis of the shaft (fig. 2 to 6). Three air motors were mounted on the carriage; one hoisted the clam, one operated the bucket opening and closing device, and the third moved the carriage along the track. The carriage was designed with guide shoes and could be hoisted to the surface in either skip compartment without trouble (fig. 4). During blasting operations the clam was tied on the rails and the carriage was hoisted up a few sets to protect it from flying rock. No attempt was made to hoist the carriage and clam together lest the clam catch in the timbers. The track on which the carriage operated was built of extra heavy 6 in. pipe with 1x2 in. strap welded on top for rails. A Yzx4 in. angle was welded on the bottom of the 6 in. pipe with a 1 in. round rod welded in the angle for a filler and additional strength. The end or cross pieces of the track which kept the track to gauge, were also made of reinforced, heavy 6 in. pipe and were bolted to the long members on which the carriage traveled (fig. 2 and 3). This section of track was used, instead of heavy railroad rails, for two reasons: (1) this section would bend before it snapped and was much safer for the men in the bottom; and (2) it was a better blasting shield, and protected the timbers better, in fact, only five wall plates were broken by blasting in 1926 ft of shaft, although 150 to 225 lb of 40 pct powder was loaded per round, depending on the ground conditions. The swinging track was supported by four % in. hoisting cables anchored to the end plates. There were also four safety cables fastened two to three sets above the carriage, which would catch the carriage if one of the supporting cables should fail for any reason (fig. 7). Clamshell Buckets: Two clams were used, each of % yard capacity, both manufactured by Blaw-Knox, Nos. 662 and 663. The No. 663 clam was 4 in. wider and 5 in. shorter than the No. 662, but seemed to dig and load more satisfactorily. It was easier to hoist and lower when it came up for repair. It did not turn over as easily, had less spill

Jan 1, 1951

By J. L. Moriarty, Irving Johnson, H. M. Feder

The crystallization of ternary alloys has been studied as a coprecipitation phenomenon by measuring the substitutional carrying of trace metals with CeCd11 precipitating from liquid cadmium. The Doerner-Hoskins logarithmic distribution law has been found to apply in all cases; the coprecipitation coefficients obtained were: 1.48 (La), 1.06 (Th), 0.63 (Pr), 0.24 (Gd), 0.17 (Sm), 0.13 (U), 0.10 (Er), 0.09 (Sr). Li, Na, K, Ba, Y, Sc, and Zr were not measurably carried. It has been observed that coprecipitation was most extensive for those tracers which form with cadmium an intermetallic compmdnd isostructural with CeCd11 or closely related thereto. For these tracers an inverse relation between Solubility in cadmium and coprecipitation coefficient was noted. THE partition of impurities between the liquid and solid phases of a metal or a congruently melting intermetallic compound has been extensively studied by the methods of zone melting.' However, when the composition of a liquid alloy and of the solid phase crystallizing from it differ grossly, partition of an impurity cannot be investigated in this manner. Partition of the impurity under equilibrium conditions may be studied by investigation of the appropriate phase diagram and partition under nonequi-librium conditions may be studied by observation of coring in the solidified alloys.2 The nonequilibrium case is of special interest in metallurgical practice because many complex alloys are cooled too rapidly to permit the uniform diffusion of impurities throughout the crystallizing phase(s). With this in mind, the present study was undertaken to obtain quantitative data on the phenomenon of coprecipitation via substitutional solid solution in certain metallic systems. It will be noticed that rather than adopting metallurgical ter- minology we have chosen to refer to the partition of impurities during alloy crystallization as coprecipitation.3 In the terminology of coprecipitation the impurity is referred to as the tracer and the precipitating phase as the carrier. From aqueous solution, coprecipitation may occur as a result of adsorption on the surface of the carrier, by occlusion of the liquid within the crystals, or by solid solution of the tracer in the carrier. By slow cooling of liquid metallic solutions it is relatively easy to form large, regular crystals of the carrier; the effects of adsorption and occlusion may be minimized and the influence of substitutional solid-solution formation may be made to predominate. Previous studies of the crystallization of some intermetallic compounds4 and of copper, silver, and gold5 have shown that the coprecipitation or partition behavior of trace metals is entirely analogous to that observed in aqueous systems when ionic substitution occurs, e.g., coprecipitation of Ra++ with BaSo4. The precipitation of CeCd11 from liquid cadmium was chosen for study. This choice was influenced by the following consideration. If the relative atomic sizes of the tracer atom and the atom for which it substitutes determine the tendency to coprecipi-tate, then a lattice having a substitutional site of well-defined size would be desirable. In CeCd116 (structure type cubic BaHg117) the cerium atom is completely surrounded by a cage of twelve nearly equidistant cadmium atoms. Thus, atoms much larger or much smaller than cerium should be poorly accommodated in this cage. The choice of a lanthanide central atom allowed the testing of the size hypothesis with a number of lanthanide tracers which differed in size but not in valence or electro-negativity. EXPERIMENTAL The general technique of this investigation was to prepare a slightly undersaturated solution of carrier in liquid cadmium to which a tracer was added. The concentration of tracer was always so low that saturation over the entire temperature range investigated was never exceeded. The solution was cooled in a stepwise fashion. After each step a por-

Jan 1, 1964

By R. M. Rose, D. P. Ferriss, J. Wulff

Single crystals of tantalum, tungsten, and the binary alloys thereof were grown by electron-beam zone melting and tested in tension between 77° and 373°K. The ductile-to-brittle transition temperature increased sharply with solute concentration, for both tungsten-rich and tantalum-rich alloys; apparently, the maximum transition temperature exists near the 50 pet concentration. Twinning was observed to be a consequence of cleavage-crack propagation. Cleavage occurred on(100)planes, and extensive tearing and "river markings" were present. The onset of brittleness could not be associated with any change in deformation mode, the yield point phenomenon, or the large rise in yield stress with decreasing temperature. Cleavage was observed to initiate at either the specimen surface, or subgrain boundaries. The observation was made that the grain boundaries of Polycrystalline materials are probably even better as nucleation sites for fracture, and are responsible for brittleness at high temperatures and lower solute concentrations than those of the corresponding single crystals. TUNGSTEN and tantalum, the two high-melting neighbors in the third transition period of the periodic table, present striking contrasts. Although both are bee, tungsten has far higher elastic moduli, and the highest ductile-to-brittle transition (DBT) of all the pure metals.' Tantalum, however, is ductile at liquid helium temperatures. Large strain aging and yield point phenomena are observed in tantalum,2,3 but only to a lesser extent in tungsten.4 Tungsten work-hardens at a high rate,5 whereas this effect is negligible in tantalum.' Since the mutual solid solubility of these two metals is complete,7 and no ordering has yet been reported in the binary system, the possibility of depressing tungsten's DBT by tantalum additions was considered. Also, tungsten additions to tantalum could produce hard, high-melting alloys which might be ductile at room temperature. Such alloys offered excellent bases for complex alloy development, particularly with regard to the effect of rhenium on Group VI metals.8"10 In addition, there was the opportunity to examine flow and fracture modes in the vicinity of the DBT, with particular emphasis on the correlation between brittleness and fundamental deformation mechanisms. Myers11 has reported that recrystallized alloys with up to 10 pet W in tantalum are ductile at room temperature, but alloys with 20 pct W could not be swaged at temperatures as high as 500°C. The whole spectrum of Ta-W alloys has been surveyed by Braun, Sedlatschek, and Kiefferl2 using sintered powder compacts. In addition to some physical properties they reported a maximum hardness at about 50 at. pet W, dropping to much lower hardness levels for the pure components. They were able to cold work tantalum-base alloys with as much as 30 pet W and to warm work tungsten-base alloys with up to 20 pet Ta by procedures similar to those used for pure tungsten. EXPERIMENTAL In order to study deformation modes and to keep the impurity concentrations at a minimum, electron-beam zone melting was used to prepare single crystals of high purity. The electron-beam zone melting apparatus is similar to that described by Calverley, Davis, and ever" and incorporates a grid control rectifier to maintain control of emission current. Pure metals were refined by making three zone passes at 4 mm per min up the rod. Alloys were prepared by melting together wire of one constituent on a previously melted rod of the major constituent. Subsequently, two zone passes were made in opposite directions to level the composition after the manner suggested by pfann.14 Purification of the starting materials was accomplished by evaporation of the several impurities, and the degree to which this was accomplished is indicated in the chemical analyses of Table I. The uniformity of composition along a zone levelled rod was within the i 1 pet accuracy of the fluorescent X-ray technique used to determine composition. Crystallographic orientation of the

Jan 1, 1962

By Albert L. Toenges

Alaska requires an adequate fuel supply for its development, and has large potential coal reserves ranging from lignite to subbituminous and anthracite. Coal production in the Territory now is less than the requirements. In L947, production was 361,000 tons, divided about equally between bituminous coal from the Matanuska field and subbituminous coal from the Nenana field. There is need for development of modern mechanized mines, which should produce the required output with a minimum of workers. These mines should not be thought of in the light of large potential capacity, as in the States. However, these new mines should be developed by modern methods, which should result in lower costs than at Present. Because of the diverse physical conditions in the coal fields, intensive investigation by diamond drilling is necessary to properly plan the development and equipment of mine sites at minimum cost. The Bureau of Mines is doing this now in the Wishbone Hill area of the Matanuska field. All coal land in Alaska is government-owned and subject to the Coal Leasing Act. Coal mining in Alaska has been handicapped by an inadequate supply of dependable workers. Wage scales have been very high in recent years, and "floaters" have been attracted to the coal mines. This type of labor is not conducive to efficient operation. Modernization of mines will require skilled and dependable workers and to induce such men to become interested in coal mining in Alaska, modern facilities, such as modern houses, schools, churches, and recreation halls, must be provided near the mines. Alaska is still a frontier and improvements in living standards must accompany development of mines in order that a stable community may be established. Coal is known to have been mined in Alaska at Port Graham, on Cook Inlet, by the Russians in 1854. During the Klondike gold rush in 1898, coal was mined in scattered areas for use on steamers plying the Yukon River and its navigable tributaries. After the Alaska Railroad was built, two important fields were developed—the Nenana and the Matanuska (Fig 1). The Bering River field was explored by the Guggenheim interests at one time, and a railroad to the area was constructed. However, the field was abandoned because of litigation and reported unsatisfactory physical condition of the beds. There are areas that today are unimportant but which may be developed in the future. Possibly the most important of these is the deposits of subbituminous coal adjacent to Cook Inlet, near Homer, on the Kenai Peninsula (see Fig 1). It is expected that the demand for coal in the Kenai Peninsula will increase because the population in that area has increased as a result of agricultural development. Numerous outcrops of coal are exposed on the peninsula and development of a mine there would obviate dependence on supplies of coal from the Matanuska field. Coal from this field must be shipped approximately 160 miles by rail to Seward and transported thence by boat about 180 miles to Homer. There has been no regular boat service from Anchorage to Homer. There are deposits of coal in the Arctic region but a description of them will not be given in this paper, which will describe the more important deposits in the Alaska Railroad belt. The results of a reconnaissance in the Arctic have been described.' Description of Fields MATANUSKA FIELD The Matanuska field lies in the valleys of the Matanuska River and its tributaries and their separating ridges. This field has been interpreted by geologists as a sunken fault block that separated the Talkeetna Range of mountains from the Chugach Range. The Matanuska River flows west along the south side of the field and the Glenn highway, which extends from Anchorage via Palmer to the Richardson highway near Glen Allan, follows the Matanuska Valley. This road to the coal field can be traveled the year around. Although the snow-

Jan 1, 1950

By Richard Hunt

BINGHAM has produced 6 per cent. of this country's copper. In total production, it ranks fourth among the copper camps of North America, the order being Butte, Michigan, Bisbee, and Bingham. In its annual production, it held that same rank until 1919, when its production was exceeded by the combined production of the Globe-Miami district of Arizona. Bingham has also been a lead camp of the first order. Utah ranks third among the lead-producing states, and its production comes largely from three camps which, in the order of their productivity, are: Bingham, Tintic, Park City. Bingham's silver production is important, but it does not rank high as a silver producer. Utah owes its rank as first among the silver-producing states to Tintic, a camp 40 miles due south of Bingham. The United States Mineral Resources reports a total production from Bingham for the period 1865-1920 having a gross value of $528,043,432. Bingham is on the eastern flank of the Oquirrh range 30 miles south-west of Salt Lake City, from which it can be reached in an hour by auto, or by the Denver & Rio Grande or the Bingham & Garfield railroads.. The automobile route is a paved road across the floor of Great Salt Lake Valley, through farming country. The Bingham & Garfield railway leaves the level of Great Salt Lake at the northern end of the Oquirrh range, and in 20 miles it climbs 2100 ft., and deposits the traveler upon a shelf cut in the side of Bingham Canyon below most of the mines, which are farther up Bingham Canyon and its tributary Carr Fork. Below lies the town, a narrow ribbon running for a couple of miles along the bottom of a canyon 2000 ft. deep-an almost impossible site for a community numbering, in prosperous times, 5000 to 7000 inhabitants drawn from all quarters of the globe. Up the canyon and down, and beneath the steep slopes nearby, and beneath the town itself, are the mining operations of over half a century.

Jan 3, 1924

By Robert H. Merrill

In 1960, the Kennecott Copper Corp. and the U.S. Bureau of Mines entered into a joint research program to determine the changes in stress, strain, and displacement created by changes in slope angles. The field site was the Kimbley Pit, located near Ely, Nev. The supporting laboratory investigations were made at the Bureau's Denver Mining Research Center. Following a study to determine the rock stresses in the mining district, the strike and dip of geologic defects (fractures) in the rock, and the laboratory-determined strength of the rock, a new slope angle and pit configuration was selected, and instruments were installed to measure changes in stress, strain, displacement, microseismic noise rate, and sonic velocities at several points in the west wall of the pit. Between February and November 1966, the west slope was changed from an average angle of 45' to a slope angle that started at 45° on the south wall, and gradually increased to 575' at the center of the west wall, and to 61° on the north wall. The pit was deepened from about 500 ft to about 550 ft. About 2.55 million tons of waste and ore were removed by stripping and mining. Measurements made between November 1965 and September 1967 showed that the pit wall was stable before excavation began, that changes in the measured quantities occurred in the pit wall during the excavation, and that the wall returned to a condition of equilibrium after the mining was completed. The need for a method and the technology to pre-design a pit slope does not require elaboration. About 5 billion tons of material are excavated yearly by surface mining methods in the United States, and the economics, safety, and availability of lower grade metals, nonmetals and fuels could all be substantially improved if operators could preselect optimum slope angles. Mining companies and the Bureau of Mines have recognized the need for developing engineering methods to determine stable slope angles. In the industry, Kennecott Copper Corp. had been making an

Jan 1, 1969

By Elliot P. Rexford

IN the early days of concrete construction, the question of quality of sand or gravel to be used as aggregate in concrete was given little or no consideration. If the construction engineer had sufficient quantity for his structure within a reasonable distance, he was contented and gave no thought to the possibility that the material might have adverse physical, chemical, or thermal properties which could cause future deterioration of his concrete and possible failure of his structure. Today a very different situation exists. With the building of larger and larger concrete structures, with a longer and longer life expectancy, the quality of the aggregate material has become of major importance. Even in small concrete structures aggregate quality cannot be passed off lightly and the quality of the aggregate must be such as to give the anticipated life without costly maintenance. The first question to arise in the selection of an aggregate is what makes a sand or gravel suitable or unsuitable for use in concrete, and how can its quality be judged. No one rock type can be depended upon to furnish sound durable material in all instances. Granitic rocks, for example, make very satisfactory aggregates in some cases, but they are among the most unsuitable of materials when they are fractured, weathered, or highly micaceous. Limestones make excellent concrete aggregates on the one hand, and on the other they may be unsound and lack durability if they contain appreciable clay, they may have a harmful chemical reaction with the cement if they contain amorphous silica, or they may have thermal characteristics which make them highly unsuitable for use in concrete. Each rock, gravel, or sand must be proved on its own merits for the particular job under consideration. The selection of suitable aggregate materials may be divided into two major phases, the geological phase which is carried on largely in the field, and the testing phase which of course is a laboratory problem; both are important. Very often the geological phase alone is sufficient for the selection of satisfactory aggregates for minor structures but for major structures both phases must be thoroughly investigated. Geological Phase In the southwest part of the country nearly every major rock type, igneous, metamorphic, and sedimentary, may be found with all imaginable intermediate phases and in all imaginable stages of alteration. Frequently very different rock types, different as to geological age, composition, and condition occur very close together. Alluvium and stream-bed sands and gravels will consist of many different rock types and even ledge rocks may vary radically as to texture and composition within the area of a quarry site. Some of the rocks make suitable aggregate materials whereas others are highly deleterious in concrete. Too often the good and the bad occur together and more than ordinary caution is required in aggregate selection. Aggregate Sources: Concrete aggregates are usually obtained from stream-bed sands and gravels or from ledge rocks. Each has certain favorable and unfavorable characteristics and it is sometimes necessary to investigate both types to determine the most suitable source for a specific structure. Stream Beds: The predominant sources of aggregate materials in southern California and Arizona are stream beds. Vast quantities of sands and gravels are brought from the mountain ranges by the rivers during times of flood but most of these streams are dry a large part of the time and hence these deposits make easily accessible sources of aggregate materials. From many standpoints stream-bed materials make very acceptable aggregates. The gravel pebbles and the sand grains are

Jan 1, 1951

By R. L. Orr, J. D. G. Masse, Ralph Hultgren

The relative partial molar enthalpies of gold and tin in liquid Au-Sn alloys have been measured from Xsn = 0.474 to 0.510 at 696°K, slightly above the melting point of solid AuSn. Within an average precision of 6 cal per g-atom, the data are represented by two smooth curves, which satisfy the Gibbs-Duhem relationship and which do not show any anomalous behavior in the range of composition studied. The results give no support to the hypothesis that atomic associations in the equiatomic liquid alloy resemble those in solid AuSu. LITTLE has been firmly established as to the nature of atomic distribution in liquid metallic phases. A favored hypothesis assumes the existence of atomic groupings in the liquid phase which are similar to those in solid phases occurring at lower temperatures. Beginning with this hypothesis, it is often possible to adjust the values of certain parameters (such as degree of association) so that approximate agreement is achieved with certain features of the X-ray diffraction pattern of the liquid, even though many other possible structures might satisfy the data equally well. Thus, in this manner, the structure of liquid tin has been described as a dispersion of fragments of gray tin in a matrix of metallic tin.1 Four X-ray diffraction studies2"5 have been made of the liquid alloy of composition AuSn at temperatures slightly above the melting point of the solid. Each investigator concluded, on somewhat different grounds, that the liquid consisted to a large extent of arrays of atoms similar to the solid, with admixture of more disordered arrays. Solid AuSn has the nickel arsenide crystal structure: each tin atom has six nearest neighbors, all gold atoms; and each gold atom has six tin and two gold atoms as nearest neighbors. It remains highly ordered to the melting point. The liquid alloy has a large exothermic (?H700° K = -2700 cal per g-atom6) enthalpy of formation, favoring ordering and making the hypothetical molecular association not unreasonable. The possible existence of such solidlike arrays, or of a high degree of order that occurs preferentially at a specific composition in a liquid alloy, might also be revealed by thermodynamic behavior somewhat analogous to that in highly ordered solid alloy phases. The latter often exist over wide composition ranges, within which the highest degree of order occurs at the ideal composition for which the atomic fractions are equal to the fractions of sites which are "right" for each kind of atom. On departing from the ideal composition, the excess atoms occupy "wrong" positions, raising the internal energy and making it progressively easier for additional atoms to occupy "wrong" positions. Thus, on either side of the ideal composition, there should be a pronounced rise in the entropy and enthalpy, which would be shown particularly by the change in the relative partial molar enthalpies, ?Hi, since they depend strongly on d?H/dx. In the one case where accurate measurements have been made, the solid AuCu ordered phase, the structure becomes much less ordered as the composition deviates from the ideal, and relative partial molar enthalpies change sharply as shown by the data6 given in Table I. This suggests the possibility of finding similar pronounced changes in the relative partial molar enthalpies of gold and tin in liquid alloys near the composition AuSn if the postulated solidlike ar-

Jan 1, 1967