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By Yu Chuang, Zai Jinmin, Liu Songyu

"The piled raft foundation is a geotechnical composite construction method, consisting of three elements, piles, raft and soils, which is applied for the foundation in an increasing number. Differential settlements have numerous negative effects on a superstructure in piled raft foundations. Through the analysis of the production mechanism of differential settlements and differential settlements control measures, a method of adjusting pile configurations is presented to control differential settlements. By comparing the measured data and finite element method analysis of one oil storage piled raft foundation, the new method is verified that the differential settlements can be reduced effectively and the inner forces of foundations are more reasonable with varying length pile configurations. The differential settlements can be minimized or even reduced to zero with right pile length.INTRODUCTIONPiled raft foundations are widely adopted for the prominent advantage of limited settlements and adequate bearing capacities. However, there always exits differential settlements in conventional piled raft foundations. The differential settlement is the main cause that produces additional stresses in the structure. Differential settlements have numerous negative effects on a superstructure as well as on a foundation, and should be restricted to allowable limits. How to control the differential settlement is the key interest.An increasing number of structures are founded on piled rafts. For this reason, design strategies that enable an optimized design have a major importance for safety and economy. An optimized design may be defined as a design achieving maximum economy of the solution, that is minimum cost for the installation of the foundation."

Jan 1, 2005

By E. M. Spokes

A.B. Rushton The work and resulting publications of the Committee on Underground Coal Mine Safety and The President's Commission on Coal were appropriate at the time of publication. The statistics cited in the publications appeared to be valid, and the actions proposed by the committee had merit. Although the actions proposed to achieve further improvements in safety may have been worthy of repeating, to associate statistics (Table 1) that are seven to eight years old with company names is bad. Statistics are readily available from the Mine Safety and Health Administration (MSHA) and are usually within three months of being up to date. A review of the most recent Coal Controlling Co. Report will reveal significant changes in the rankings. The more recent statistics should have been cited, or at least compared with those used in the article. Westmoreland Coal Co. is listed as having a 21.1 disabling injury rate for the period (1/78-12/80). The disabling injury rate for the Company for the year ending 1985 was 7.4 for the underground mines and 6.25 for the total facilities. Through March 1986, the disabling rate for the underground mines was 2.86. The management and employees of Westmoreland Coal Co. are committed to safety. The improvements achieved over the past five years attest to that commitment. We are justifiably proud of the achievement and dislike being misrepresented as in the April 1986 article. Reply by E.M. Spokes I regret that the lapse of time since the original work gives a picture of safety practice that is no longer true today. That Westmoreland has reduced its disabling injury rate by 65% to 7.4 in 1985 is most commendable, and I congratulate you for this accomplishment. It is typical of the progress sought by the National Academy of Science and National Research Council when they funded the original study and published "Toward Safer Underground Coal Mines" in 1982. If a significant number of companies do as well as Westmoreland, the report will have served its purpose. Since the paper was primarily a condensation of the original report, it would not have been appropriate to seek access to later statistics. The purpose in seeking its publication at the time of presentation was to reach those who might have missed the original report. The fact that publication was delayed has not voided that purpose and may have reminded some who were aware of the original that its recommendations still apply to improvement of underground coal mine safety.

Jan 6, 1986

By Guangyuan Li

A 3-D block-spring model (BSM) has been developed for simulating the behaviour of jointed rocks. The model uses a discontinuous formulation and an iterative procedure to solve the large displacements in jointed and faulted rock masses. The 3-D BSM has been used to simulate mining excavations in jointed rocks around an open pit and underground settings. The applications have demonstrated that it is a practical tool to solve complicated problems of rock mechanics in civil and mining engineering. This model has recently been modified to incorporate complex geological conditions, such as underground water flow (seepage), unstable blocks and the time-dependent behavior of backfill materials. To demonstrate the capability of the 3-D model this paper presents the fundamentals of the 3- D BSM and its application for open-pit mining. A complete computer package BSM3D has been developed based on the 3-D block-spring model. The package consists of a solving engine for numerical analysis and a Windows based processor for mesh generation and results display. The multi-functioned pre-processor provides tools and menus to help users setting up models easily and conveniently. The fully featured post-processor allows users to view the results in different windows and orientations. The graphical results include contours, vectors, animation of rock movement, and the failure state of the rock structures.

May 1, 2001

By Pei-Yung Huang, Carle R. Hayward

IN order to understand more fully the actual state of copper lost in copper reverberatory slags, a systematic study on the various reactions between certain metallurgically important copper compounds and synthetic copper reverberatory slags was undertaken by the authors in the Massachusetts Institute of Technology from 1942 to 1945.1 This paper, being only a part of the work completed, is published first because it furnishes the first experimental evidence of the formation of copper ferrite from reactions between cuprous oxide and copper reverberatory slags. CHOICE OF SLAG COMPOSITION The choice of the composition of the synthetic reverberatory slags to be used was much deliberated. Commercial copper reverberatory slags, neglecting minor constituents, average:2.3 45 to 47 pct FeO; 38 to 42 pct Si02; 5 to 6 pct CaO; 5 to 6 pct A1203. Since the ternary system CaO-FeO-SiO2 has been much more thoroughly studied4 and is much simpler than the quaternary system CaO-FeO-Si02-A1203, it was decided to use calcium oxide only instead of calcium oxide and alumina in the synthesis of the slags. Fortunately the composition of the average copper reverberatory slag with calcium oxide substituted for alumina lies in the region of the phase diagram clarified by the investigators. A survey of the phase diagram (Fig I) suggested two compositions: (I) the com¬position near the invariant point S, which is made up of 46 pct ferrous oxide, 37 pct silica and 17 pct calcium oxide; (2) the composition near the invariant point K with 45.5 pct ferrous oxide, 43 pct silica and 11.5 pct calcium oxide. Composition S has the advantage that the percentages of ferrous oxide and of silica almost exactly coincide with those of the average reverberatory slags, and that of the lowest melting point (1093° ± 3°C) in the whole ternary system. The calcium oxide content, however, obviously is high. Since we have already substituted calcium oxide for alumina, and since calcium oxide is more basic than alumina, the resulting slag is decidedly too basic. Composition K has a slightly higher melting point (1105° ± 3°C) and a higher silica content, but this is preferred because the slag is saturated with respect to tridymite and thus should not corrode siliceous refractories. Besides, it is less basic than composition S. The final slag composition, therefore, was chosen to

Jan 1, 1947

By Robert D. Harris

In 1958, the coal industry, through Bituminous Coal Research, Inc., and the electric utility industry, through the Edison Electric Institute and the Association of Edison Illuminating Companies, initiated a joint program to study coal pyrite and its removal as a means of lowering the sulfur content of steam coals. It was assumed that significant amounts of pyrite often remained in washed steam coal not because available systems for coal cleaning were inadequate, but because the pyrite was small and widely disseminated, During the first two years, samples of coal were examined petrographically to determine the size, association, and mode of occurrence of the pyrite in the raw coal and in the products of laboratory separation, Samples of coal were also pulverized in different equipment to determine how effectively pyrite was released by pulverization and if the mode of pulverization affected the liberation of pyrite. From these studies, the following tentative conclusions were made (1 ): 1. Much of the coal pyrite is within a 250-70 micron size range, and therefore is too small to be removed by conventional coal cleaning, but relatively large when compared to the size of coal as pulverized for combustion in large boilers. 2. The degree of pyrite liberation is mainly determined by the size of the product and not by the type of crusher. 3. For improved pyrite removal, the coal should be cleaned during or after pulverization at the power station.

Jan 1, 1965

By Frank B. Fulkerson

Three of the five primary aluminum producers in the Pacific Northwest have facilities in the region to make aluminum mill products. Only one of the three, however, produces a variety of shapes and forms. The principal reason for the lag in development of an aluminum mill products industry was found to be transportation cost. The long range outlook is for expansion, based mainly on growth of the western market for aluminum used in construction, containers and packaging, electrical applications, diesel trucks, mobile homes, and irregation pipe. Consumption of western aluminum mill products in 1985 is projected at 1.1 million tons, compared with estimated consumption of 250,000 tons in 1961.

Jan 1, 1965

By G. Sachs, G. Espey

It is well known that high residual stresses are created in tubing by the sinking process, in which no internal tool or mandrel is used.l-4 In this process, the wall thickness is usually slightly increased.4,5 In tube drawing with a plug, the residual stresses are gradually reduced with increasing reduction of wall, and, by the conventional approximation methods of stress measurement, no stress is detected in tubing in which the wall reduction is over 50 per cent of the total reduction in cross-sectional area.l,6 It has also been observed that higher stresses are set up by the use of wide-angle dies than by acute dies.? 'he actual stress distribution in drawn tubing has been investigated by a few workers,7"9 but has not been related to the conditions of the drawing process. Some attempts have been made to correlate accelerated cracking tests for stresses in brass products with the magnitude of the circumferential stresses and the condition of the metal. l-10 Materials and Preparation The material used in this experiment was commercial cartridge-brass (70 per cent Cu, 30 per cent Zn) tubing in the straightened and annealed condition.† This tubing was of various outside diameters and had a wall thickness of 0.032 in. It was sunk to 1/2-in. outside diameter by means of an experimental 10,000-lb. hydraulic draw-bench, with a maximum drawing speed of 8 ft. per minute. All dies were of 1/2-in. diameter. Various die materials and contours were used in order to study the effect of die design on residual stresses in the tubing; the specifications for the dies are given in Fig. I. Castor oil was used throughout as lubricant. Chattering occurred in sinking some of the tubing, particularly when dies having the more acute angles were used, but the tubes showing chatter marks did not differ markedly from those with a smooth surface regarding the magnitude of residual stress. The source of the chattering has not yet been revealed. The wall of the tubing becomes thicker during s nking, as previously observed.4,5 The increases in wall thickness were 3, 7 and 9.5 per cent on the average for 6, 20 and 33 per cent reductions in diameter, no effect of the contour being noticeable. The reductions mentioned throughout this investigation do not consider this fact, but are reductions of the outside diameter. For the evaluation of the residual stress two fundamentally different methods were used. The approximation methodl,ll of splitting a certain length of tubing was employed to obtain a bird's-eye view of the effect of various drawing conditions on the circumferential stress measured by the diameter change occurring On splitting. For

Jan 1, 1942

By G. Sachs, G. Espey

It is well known that high residual stresses are created in tubing by the sinking process, in which no internal tool or mandrel is used.l-4 In this process, the wall thickness is usually slightly increased.4,5 In tube drawing with a plug, the residual stresses are gradually reduced with increasing reduction of wall, and, by the conventional approximation methods of stress measurement, no stress is detected in tubing in which the wall reduction is over 50 per cent of the total reduction in cross-sectional area.l,6 It has also been observed that higher stresses are set up by the use of wide-angle dies than by acute dies.? 'he actual stress distribution in drawn tubing has been investigated by a few workers,7"9 but has not been related to the conditions of the drawing process. Some attempts have been made to correlate accelerated cracking tests for stresses in brass products with the magnitude of the circumferential stresses and the condition of the metal. l-10 Materials and Preparation The material used in this experiment was commercial cartridge-brass (70 per cent Cu, 30 per cent Zn) tubing in the straightened and annealed condition.† This tubing was of various outside diameters and had a wall thickness of 0.032 in. It was sunk to 1/2-in. outside diameter by means of an experimental 10,000-lb. hydraulic draw-bench, with a maximum drawing speed of 8 ft. per minute. All dies were of 1/2-in. diameter. Various die materials and contours were used in order to study the effect of die design on residual stresses in the tubing; the specifications for the dies are given in Fig. I. Castor oil was used throughout as lubricant. Chattering occurred in sinking some of the tubing, particularly when dies having the more acute angles were used, but the tubes showing chatter marks did not differ markedly from those with a smooth surface regarding the magnitude of residual stress. The source of the chattering has not yet been revealed. The wall of the tubing becomes thicker during s nking, as previously observed.4,5 The increases in wall thickness were 3, 7 and 9.5 per cent on the average for 6, 20 and 33 per cent reductions in diameter, no effect of the contour being noticeable. The reductions mentioned throughout this investigation do not consider this fact, but are reductions of the outside diameter. For the evaluation of the residual stress two fundamentally different methods were used. The approximation methodl,ll of splitting a certain length of tubing was employed to obtain a bird's-eye view of the effect of various drawing conditions on the circumferential stress measured by the diameter change occurring On splitting. For

Jan 1, 1942

By J. B. Gayle

Coke-pushing difficulties have been the subject of much controversy but relatively little scientific investigation. Nevertheless, continuing pushing difficulties can greatly decrease the life of a coke oven and should therefore be considered in the selection of the blend and operating conditions used for coke production. The present Federal Bureau of Mines investigation was undertaken to correlate results of previous investigators and to supplement these results to the limited extent practical by us of an experimental slot-type oven. To this end, pushing pressures have been determined routinely for Tuscaloosa oven tests conducted during the past 10 years. Nevertheless, the accumulated information remains incomplete in many respects, and in view of the recognized inadequacies of the data, publication is now undertaken only because of the scarcity of published information on this important problem and the discontinuation of carbonization studies with this oven.

Jan 1, 1961

By Gerald Verbeek, Erick Baziw

Downhole Seismic Testing (DST) is an extensively utilized geotechnical tool for site characterization. DST provides low strain (<10-5) in-situ interval shear and compression wave velocity estimates. These velocities are determined by obtaining relative arrival times of source waves as they travel through the stratigraphy and are recorded by one or more vertically offset seismic sensors. The DST arrival times have associated measurements errors and resolution limitations, which become more pronounced as the depth interval of analysis is reduced. Baziw Consulting Engineers (BCE) has developed a new DST analysis technique, the so-called DSTPolyKF algorithm, whereby the arrival times of DST data set are used to define a high order polynomial. The main advantages of this new technique are four-fold: 1) the ability to utilize all arrival time estimates irrespective of measurement errors. 2) the ability to analyze the data with user specified and even small (≤ 0.5m) depth intervals. 3) the ability to apply sophisticated data fusion for significantly more accurate DST interval velocity estimation. 4) the use of polynomial regression accuracy parameters to quantify how well the “best fit” polynomial fits the acquired arrival time data sets. This paper outlines the mathematical details of the best fit polynomial where a Kalman filter formulation is implemented. The performance of the DSTPolyKF algorithm is demonstrated by processing real DST data sets.

Oct 1, 2022

By W A. Peck

Raise boring is the preferred option for constructing ventilation shafts in Australia. The trend has been towards larger diameters and longer single-pass shafts, and if at all possible, reaming through to the surface. The risk of raise failures, due to missed, unforseen or latent ground conditions has therefore been increasing, along with the need for better, more confident stability assessments.While the stability of specific weak zones and structures must be assessed individually, empirical rock mass classification schemes have been successfully used for about the last 20 years to assess general raise stability; eg the McCracken and Stacey method (M&S).Unfortunately M&S, as originally documented in McCracken and Stacey (1989), did not include a database of raise diameter and rock mass conditions versus actual performance. Nor was there a procedure for assessing the risk of failure due to adverse ground conditions. Peck and Lee (2008) published an Australian database of raise bore performance. Raise diameter and ground conditions, in terms of a lower bound Qr value, were presented for 50 cases versus raise performance. This paper updates the 2008 raise performance plot with more and SRF-corrected data. McCracken and Stacey (1989) had previously recommended that additional geotechnical data should be acquired and analysed when the proposed raise plotted on the unstable side of raisebore stability ratio (RSR) = 1.3, but the Australian data used in this study also suggests that all sites with Qr Contributions and sensitivities of various rock mass parameters (RQD, Jn, Ja, Jr, Jw, SRF, time-dependent weathering/deterioration) to raise stability and risk of failure are discussed. Numerous case studies are used to illustrate issues with stability assessments.As raises can now be remotely supported using sprayed fibrecrete, questions of time-dependency are also discussed. In recent times, contiguous piles, ‘mini-piles’ and grouting have also been successfully used to presupport or otherwise improve the stability of near-surface sections of weathered, or poor quality rock masses.

Mar 21, 2011

By Summers D. A.

A mining machine has been designed and fabricated to mine coal from room and pillar operations combining high pressure waterjet cutting and mechanical fragmentation. Waterjets are used to slot the coal face with an advancing plow wedge then breaking out larger coal pieces. The machine has the following advantages: - simply built around a disassembled longwall chock and gathering arm loader apron; - few moving parts which are simply and inexpensively maintained; provides immediate roof support in the face working area; - coal produced is larger, with resulting lower preparation cost; - no dust is generated; - risk of spark generation is virtually eliminated.

Jan 1, 1992

By S. Pollak, J. Rickard

The Long Baseline Neutrino Facility Far Site in Lead, South Dakota is currently under construction and involves drill and blast excavation of the largest and deepest caverns in North America. This paper will review how the geotechnical risks are being managed through careful consideration of excavation sequence, completion of exploratory pilot tunnels, implementation of a robust instrumentation and monitoring program, and validation against the design. Aside from having a cross section of 5,900 square feet and being situated at a depth of one mile, the caverns also employ a unique permanent rock bolt solution which underwent a significant pre-production pull testing program involving varying installation methods and being subject to blast vibrations. The behavior of the schistose rock mass and performance of the ground support are being continually monitored throughout the construction phase.

Jun 13, 2023

By W. H. Ode

Because trade increased among various nations after World War II, the Coal Committee of the Economic Commission for Europe established a Classification Working Party in 1949 to develop an International system for classifying coal. Although national classification systems existed, they had been developed in accordance with characteristics of domestic coals without reference to coals of other countries. Terms for describing similar or identical coals differed in each nation. This caused confusion in evaluating coals shipped in international trade and also hampered exchange of technical information on problems of coal utilization. Experts4/ participating in the classification project early in their work agreed that two systems should be formulated, one for the higher rank coals or so-called "hard" coals in European terminology and the other one for the lower rank coals, the latter designated mainly as brown coals and lignites in the various countries. Rank refers to the degree of metamorphism (progressive alteration) in the natural coalification series from lignite and brown coal to anthracite. The system for classification of hard coals was published late in 1956 as a document of the United Nations.5/ Results of a study for applying this system to American coals were published in 1958.6/

Jan 1, 1960

By C. K. Man

Coal dust explosions in underground coal mines are prevented by generous application of rock dust (usually limestone). If an explosion should occur, the rock dust disperses, mixes with the coal dust and prevents flame propagation by acting as a thermal inhibitor or heat sink. To investigate this process in more detail, a number of explosion experiments using various coal dust and limestone rock dust mixes have been carried out at the Lake Lynn Experimental Mine (LLEM). These were conducted in a single entry of about 1600 ft (488m) in length and initiated with a methane gas explosion. A consistent set of post-explosion floor dust samples were taken along the entry after each test. These dust samples have been analyzed in the laboratory using thermogravimetric analysis (TGA) and solubility to determine how the limestone rock dust behaved during the coal dust / rock dust explosions. The preliminary results are reported in this paper and indicate that the chemistry of limestone plays an important role in its capacity to inhibit coal dust explosions. The results also show that it may be possible to estimate the intensity of the explosion using these conventional methods of analysis. Disclaimer: The findings and conclusions in this report have not been formally disseminated by the National Institute for Occupational Safety and Health and should not be construed to represent any agency determination or policy.

Jan 1, 2009

The Northparkes E26 Mine is AustraliaÆs first block cave mine. This paper presents an outline of the design and construction philosophies that drove the project development and details the excavation, ground support and construction methods used. Features of the block caving method at Northparkes include:high production rates from multiple drawpoints;hard blocky ore fractured along gypsum veins; andloading/unloading of openings due to mining induced stresses.The drawpoints are subjected to high abrasion from the flow of ore, as well as impact from both secondary blasting and from the production LHDs. In order to withstand these harsh conditions and to ensure a low maintenance environment, a conservative approach to ground support was taken. The experiences gained during construction are discussed and whether the design expectations have been achieved. By installing a high level of ground support during excavation of the openings subject to mining stresses, a low maintenance environment has been constructed. Operating costs are therefore minimised and Northparkes E26 Mine continues to produce at the lowest cost per tonne of all Australian underground mines.

Jan 1, 1999

By H. U. Ross

ROASTING is usually thought of as a process in which finely crushed or ground ore is heated with air in a suitable furnace in which the ore is rabbled or agitated in such a way as to insure complete reaction between the oxygen of the air and the oxidisable constituents of the ore. Roasting in this sense is used only to a very limited extent in the iron and steel industry. This is because, in most iron ore deposits mined today, the ore is of a type that does not require roasting. Where this process is used, it is usually because of some special circumstance, such as a peculiar economic condition, that warrants it. If roasting is looked upon in the broader sense as a gas-solid chemical reaction carried out at an elevated temperature in which both gaseous and solid products are formed, then the process is used in the iron and steel industry to a greater extent

Jan 1, 1961

By Walter I. Nelson

THE geology of the Copper Mountain mine has been ably described by Dr. Victor Dolmage in Memoir 171 of the Geological Survey. Reference is invited to that paper for a more extended description of the geology of the ore deposits than will be given here. The important orebodies of the mine occur in fragmencal volcanic rocks, parallel with and adjacent to the gabbro of the Copper Mountain stock. In the mine area, the contact strikes N.45°W. and dips either vertical or steeply co the southwest. For a distance of 500 feet or more from the contact, the fragmental rocks have been intensely metamorphosed both dynamically and hydrothermally. The dynamic effects are reflected in the foliation and fracturing parallel with the contact, which occurs throughout the mine, and the volcanic rocks within this foliated zone are, in places, highly schistose. Superimposed upon thi.s structure is a system of fine, straight, parallel frac-tures that strike nearly normal to the foliation (and contact) and dip between 65° and 75° to the north; it is with this last-described fracture system that the ore minerals are structurally associated. The fractures are mineralized with bornite and, to a considerably less extent, with chalcopyrite and some pyrite. A notable characteristic is the extreme fineness of the fractures; even in high-grade ore they are rarely larger than pencil lines. The tenor of the ore, therefore, varies directly with the fracture spacing. Replacemcnr of the host-rock between two or more fractures, varying in degree from finely disseminated sulphides to masses of nearly solid bornite up to 18 inches wide and 2 to 3 feet long, is not uncommon in the higher-grade portions of the orebodies. Two structural types of orebody are recognized in local nomenclature: the &apos;contact&apos; orebodies, so-called because they occur along the contact; and &apos;outlying&apos; orebodics, which are enclosed in the volcanics. The former arc tabular or lenticular bodies elongated along the strike and dip of the contact. The &apos;outlying&apos; orebodies are coniform, chimney-like masses roughly elliptical in plan; with depth, their roots plunge at steep angles toward the contact. All of the orebodies are cut by a number of light-coloured granophyre and felsite porphyry dykes. These dykes-the so-called &apos;mine&apos; dykes -enter the mine area from the north, curve gently southeasterly across the orebodies, and then cross the contact into the gabbro at the south end of the mine. All the dykes dip vertical to steeply east.

Jan 1, 1941

By Donna L. Anderson

This U.S. Bureau of Mines report describes a system to determine the position and heading of a continuous mining machine during the maneuvers required at the face area of a mine. The system consists of commercially available angular position sensors and a computer interface developed and tested at the Bureau. The sensors employed are laser-based scanning devices that report the angular location of retroreflective targets within their 90° field-of-view. The targets are mounted on the continuous mining machine; the laser scanners are mounted on an external reference structure. The computer interface consists of a single-board computer-controller programmed to gather and process the angular position information, and to calculate the machine&apos;s position and heading relative to the reference structure. A sensor fusion algorithm will be incorporated into the computer interface programming to increase the accuracy and reliability of the system.

Jan 1, 1989

By W. C. Ellis, M. E. Fine

INTRODUCTION IN the iron-cobalt system there are several property-composition relationships of theoretical importance. The alloys are ferromagnetic exhibiting a maximum saturation at approximately 33 at. pct cobalt;1 the interatomic distances at room temperature show large positive deviations from Vegard&apos;s rule;2 and in the region of 50 at. pct cobalt there is an order-disorder transition which has a large effect on other properties. The thermal expansion data for the system have been reported3,4.5,6 only from room temperature to 100°C, and for this temperature range a minimum in coefficient in the region of 50 at. pct cobalt is well established. 3,4,6 This paper reports the expansion properties in the iron-cobalt system for temperatures from 30 to 850°C, shows the correlation with other physical properties, and proposes an explanation for the interrelation. EXPERIMENTAL PROCEDURE The thermal expansivities of iron-cobalt alloys in 10 pct steps in composition were determined from 30 to about 850°C. The major series of alloys was prepared by induction melting from Armco iron and cobalt rondelles with the addition of about 0.6 wt pct manganese. This series was supplemented by three compositions: high purity electrolytic iron, high purity electrolytic cobalt, and an alloy of 52.1 wt pct cobalt prepared from the high purity electrolytic iron and electrolytic cobalt by vacuum-melting. The compositions of both series of alloys are given in Table I. The specimens for measurement were prepared from the cast ingots by first hot-swaging followed in most instances by cold-swaging to 0.180-in. diam rod. Some alloys in the region of 50 pct cobalt, because of inherent cold-brittleness, were swaged hot to the final diameter. The expansion specimens were approximately 2 ½ in. long with flat-ground and polished ends. Before measurement they were annealed in hydrogen for r hr at 900°C. In the apparatus used for the thermal expansivity measurements (Fig I), a modification of that designed by Hidnert and Sweeney,7 the differential expansion between the specimen and vitreous silica was measured by a dial micrometer reading to 0.001 mm. An invar fixture supported the dial gauge, and this fixture, in turn, was connected to a split invar collar which supported the outer silica tube. A pyrex "T" was placed over the outer silica tube and was sealed to the invar collar by Apiezon Q sealing compound. A movable hollow vitreous silica rod which rested on top of the test specimen slid through a flexible rubber diaphragm cemented to the top of the pyrex "T." The specimen rested on a vitreous silica pedestal fused into the outer tube and was kept centered by a silica sleeve. A non-oxidizing atmosphere was obtained by admitting hydrogen into the

Jan 1, 1948