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By John B. Goddard

In-situ leaching of uranium by aqueous ammonium carbonate containing oxygen or hydrogen peroxide as oxidant results in the partial dissolution of sulfides. While some of the sulfide sulfur is oxidized to sulfate, a considerable portion is oxidized only to thiosulfate and polythionates. Polythionates poison the anion exchange resin used to extract uranium. At the Palangana operation in southeast Texas, the trithionate ion (S3062-) was found by infrared spectrometry to be the major resin poison. Treatment of ion exchange feed solutions with hydrogen peroxide only resulted in partial conversion of reduced sulfur species to sulfate, and trithionate conversion was particularly slow. Poisoning of the resin also occurred from the ammonium chloride eluant, which built up trithionate because of recycling. This was essentially eliminated by chlorinating the eluant prior to yellow cake precipitation.

Jan 1, 1984

By Noel Cunningham

MANY excellent descriptions of the mills of the, Porcupine district have been written, but no discussion exclusively devoted to the metallurgical technology has been given. These notes are intended to cover this feature briefly. They are based upon 2 ½ years' mill operation in the district-i.e., practically since the beginning of metallurgical operations. Character of the Porcupine Ore.-There is no oxidized ore in the district, the surface having been deeply planed by glacial action in recent geologic time. The precious-metal content is about in the proportion of 85 of gold to 15? of silver by weight; hence, the silver is practically negligible. There are two classes of Porcupine ore, having very different characteristics; these will be referred to throughout this paper as Class A and Class B. Class A ore is a pure quartz with inclusions of schist. Generally it is heavily fractured .and breaks down readily to sharp, hard grains, about minus 10 plus 20 mesh, requiring further comminution to release the gold. It carries very little pyrite; the gold is entirely free and apt to be coarse, but often spongy, going into solution readily on that account. This gold is 60 per cent. to 85 per cent. free milling, depending on the grade of ore. Class B ore is an iron silicate schist, strongly laminated, carrying 4 to 5 per cent. pyrite; its specific gravity is 2.8 to 3.0, depending upon the amount of mineralization. In breaking the ore in the mine, generally over 25 per cent. of material through a ½ -in. ring is made; the ore readily breaks down in milling and makes a comparatively large amount of non-crystalline slime; owing to its high specific gravity, however, it is quick settling.' In my opinion, the gold in this ore is free, but so finely divided that it will neither pan nor amalgamate; it appears to be disseminated through the rock and not chiefly associated with the pyrite. Veins of Class A ore occur with or without side walls of Class B, and veins of Class B occur unassociated with Class A; more often the veins are closely banded, Class A and Class B alternating, generally with Class B in excess. Both classes of ore are more-or less blocky at times, and with ref¬erence to Class B this is indicative or low gold content.

Jan 3, 1915

By L. E. Young

MUCH has been said and written on this subject and probably little new can be said. However, the point of view of the operating company changes from time to time, and more stress may be laid upon a subject today than was done some years ago, and it is with this changing point of view that the following observations are presented: The writer wishes to limit the discussion to views of a company mining bituminous coal, particularly to what the production department expects of the college grad¬uate. In an editorial in MINING AND METALLURGY for February, 1929, entitled "Is. Mining Engineering Passing?" it is stated that: It is not by accident that mining engineering courses have been general rather than special and form even today the only general engineering courses widely offered and followed by engineering students.'* * * even in this age of specialization, the mining engineers remain a body characterized particularly by broad knowledge.

Jan 1, 1929

By Kenji Monden

"The creep properties of tin-based lead-free solders Sn3.0AgO.5Cu and SnO.7CuO.1Ni and lead solders 10Sn90Pb and 50Sn50Pb were investigated at temperatures between 298 K and 398 K. The creep-rupture time decreases with increasing initial stress and temperature. The analysis of the solder creep curves uses the Omega method. The creep rate E is expressed by following formula: 1n ? = 1n ?0 + O?, where ?0 and O are experimentally determined. The parameter Eo increases with increasing initial stress and temperature. The activation energy for ?0 is 108 kJ/mol for Sn3.0Ag0.5Cu and 85 kJ/mol for SnO.7CuO.INi. These energies suggest that the lattice diffusion of tin is dominant. The energy of 1OSn90Pb is 37 kJ/mol and that of 50Sn50Pb is 67 kJ/mol. The creep-rupture time is calculated using the parameters, ?0 and O. The calculated creep-rupture time is in good agreement with the measured creep-rupture time.IntroductionLead-free solders are replacing traditional tin-lead solders as part of a major trend toward protecting the earth's environment. However, tin-lead solders are currently needed for a high temperature range. Since creep is the main deformation mechanism in solder joints, understanding a constitutive model for creep and examining creep data for solders are important for reliability analysis of solder joints. A steady-state creep equation is generally used for solders [1,2], and it supposes that the secondary-, or minimum-, creep rate occupies the entire creep life. However, secondary creep accounts for only about 15 % to 40 % of the entire creep life for solders. Thus, the equation may not represent the entire creep life."

Jan 1, 2012

By Joy Sandberg

The 1989 mining industry compensation survey done annually by the Denver-based Mountain States Employers Council shows the "average" or "typical" employee in the mining industry is projected to receive an increase of 3.9% to base pay in 1990, compared to a projection of 4.1 % for 1989. According to MSEC, the actual percentage increase to base pay has continued to decline since 1980. A review of pay increases since 1980 shows the highest average increase was 11%, granted in 1981. Since 1981, there has been a steady deceleration through the decade. While 1989 final figures are not available, 1988 numbers might indicate a reversal of this trend. In 1988, the average employee actual increase was 3.8%. That was 0.7% higher than the 3.1% projection. This may be significant as actual increases have been lower than projections since 1982. Chart A shows the pattern of projected vs. actual percent increases in pay since 1980.

Jan 1, 1990

Caving of the second lift of the E26 orebody at Northparkes Mines was monitored with a comprehensive package of instrumentation, including an ISS seismic monitoring system. Seismicity was collected by an array of 19 triaxial accelerometers and 9 uniaxial geophones, giving a high quality seismic data set. The seismic response to block caving at Northparkes has been divided into two papers. The first paper focuses on the seismicity during undercutting. The second paper deals primarily with the seismicity during cave propagation during initial cave production. In this paper, traditional seismic monitoring concepts are applied to seismicity during cave mining. More than 40 000 seismic events were recorded during cave undercutting between February and January 2004. Events up to local magnitude +1.3 were recorded near the mine during undercutting, with event rates of up to 500 per day. Analysis of frequency-magnitude distributions, S-wave to P-wave energy, and temporal variations in event magnitude all gave insight into the seismicity recorded. In addition, a new technique, called apparent stress time history, was shown to be sensitive to stress change from undercut blasting, and useful to identify significant periods during cave development. A significant difference was identified between the cave back location and the halo of seismic events above the cave. Throughout the undercutting, the a seismic zone of loosening was typically 50 to 70 metres in vertical thickness. Based on the seismicity, a stress driven rock mass failure model is proposed. Early in the undercutting, seismicity and movements in the seismogenic zone were directly triggered by stress change from undercutting. Towards the end of undercutting, seismicity was virtually uncorrelated with undercut blasting, and likely due to the onset of cave initiation.

Jan 1, 2008

By D. H. Garcia

The mining industry responded to public concerns about use of cyanide by developing a voluntary certification to allow companies to demonstrate their commitment. The International Cyanide Code (Code) establishes standards for the manufacture, transportation and use of cyanide. Key concepts in closure planning to comply with the Code are to decommission all facilities related to cyanide operations and to provide a financial assurance mechanism for those activities. Code requirements may differ from state-specific regulations, but requirements are based on closure goals for protection of the human health and the environment. Only issues related to cyanide operations are required to be included in decommissioning demonstrations. For example, the disposal method of cyanide solution holding tanks and the cost must be included in the closure plan, but information regarding closure of roads is not required. Closure costs must be based on a third-party contractor as opposed to internal budgets. Compliance with the closure requirements of the Code allows companies to show their stockholders, lending institutions, and the public that their operations are environmentally responsible.

Jan 1, 2010

By G. A. Kordosky

LIX® 34 liquid ion exchange reagent, one of a totally new class of metal extractants, is discussed in terms of its properties. Specifically, these properties include improved selectivity for copper and greater recovery. Some comparisons to the well known General Mills' LIX® 64N have been made. Particular emphasis is given to treating solutions containing very low copper values and high impurity levels.

Jan 1, 1976

By G. M. Furnival

A PROGRAMME of structure drilling was commenced in the Southern Alberta Plains in June, 1943, for the purpose of determining the geological structure of the underlying sediments. More than 270 holes have been drilled throughout an area of some 1,800 square miles or 1,000,000 acres (see Figure 1). The Southern Plains of Alberta are covered by a mantle of glacial drift,' at places exceeding 200 feet in thickness. Only where this mantle has been dissected by deep drainage channels is bed-rock exposed to any great extent. Along such channels, surface geological mapping can be conducted to ad- vantage. Accordingly, geological mapping was commenced _in 1942 and continued each season to the present. To supplement this information, a survey of the water wells was carried on in conjunction with the mapping. However, to obtain more positive and evenly distributed structural control throughout the area as a whole and to define more adequately structural trends indicated by the preliminary mapping, it was necessary to resort to the use of drills to penetrate the glacial mantle. Various types of light drills are commonly used for this purpose and provide geological information in the form of drill cuttings, cores, drilling time rates, etc. Correlation of information is generally obtained through comparison of cored sections, drill cuttings, micro-fossils, and drilling time rates. However, in this case the holes were electrologged and correlations were made chiefly by comparison of such logs. The data so obtained is plotted and portrayed in the form of structure contour maps on recognizable stratigraphic horizons and in the form of isopach maps of the recognizable stratigraphic intervals penetrated in the holes. This method had not been widely used in Alberta previously and the technique employed was necessarily developed after the project began.

Jan 1, 1945

By A. F. Brock

Introduction During the past ten years, shaft sinking has been going on continuously at Inco and new shafts and shaft extensions have been sunk under a broad diversity of conditions. The Murray mine and the Stobie section of the Frood-Stobie mine have?been opened and a main hoisting and supply shaft has been sunk at each of these mines; No. 2 shaft at Murray was sunk initially from surface to 1,960 feet and is now being extended; No. 7 shaft at Stobie was sunk from surface to 3,105 feet. No. 2 shaft at Garson has been extended from 2,200 feet to 3,200 feet in depth. At Levack, No. 2 shaft is being extended from 1,860 feet to 2,850 feet for ore hoisting from the 2,650 lev el. At Creighton, No. 6 shaft, collared on 52 level, was sunk 1,740 feet to a total depth of 5,563 feet below surface. It is the deepest Inco working. At Creighton, also, No. 7 shaft is being sunk from surface to the 30 level to handle the ore from caving operations. Ail Inco main hoisting and supply shafts sunk during the past twenty years have been vertical rectangular shafts equipped with two cages and two ore skips. The cages are standard and interchangeable for all shafts except Garson No. 2; they are single-deck, 13 ft. 4 in. by 5 ft. 2 in. inside and are designed to carry a maximum material load of 16,000 lb. The man load is 48 men. Skip capacities at the several shafts vary from 8 to 15 tons. Forty-two hundred feet bas been established as the standard vertical lift, and the permanent hoisting equipment at each shaft is designed for eventual hoisting from that depth.

Jan 1, 1950

By Edward Thomas

In 1947 the Bureau of Mines, as an accident-prevention measure, decided to advocate the adoption of roof bolting (suspension roof supports) in coal mines where it appeared to be applicable. Subsequent experience has disclosed that the applicability of this support is more general than was first suspected; moreover, comparison of the accident records of mines before and after adoption of the roof-bolting method shows results that more than justify the effort and funds expended. The industry has responded enthusiastically, because roof bolting proved to be not only an accident-prevention measure but also one to increase productive efficiency. In some instances this increase in the productivity of men and machines has been almost incredible and has led many mining men to conclude that through some hocus-pocus they can now cast aside many or all of the time-proven rule of good mining practice. If this tendency is not counteracted, it could easily result in an increase of roof-fall accidents, which may outweigh the benefits of roof bolting and discredit an innovation in roof control that shows great promise of materially reducing our mining accidents. From an economic standpoint it may represent the margin between profit and loss in many mining operations. The purpose of this manuscript is not to propose an inflexible code or set of rules but to summarize the experience to date as a guide for those who have the responsibility for the installation and inspection of roof bolts. This experience consists of approximately 250 investigations of roof failure in which roof bolting was involved either directly or indirectly. It should be pointed out that many of these failures were controlled, occurring during experimental periods, and comparatively few of the accidental failures actually resulted in personal injury, yet a significant number represented potential hazards.

Jan 1, 1951

By S Hapugoda, M Peterson, A Haileslassie, A Poliakov, E Donskoi, L Lu

Most metallic iron in the world is produced in blast furnaces from different feed stocks – sinter, lump ore and pellets. Sinter is the main iron-containing feed to the blast furnace and can constitute up to 70–85 per cent of the total ferrous burden. As sinter quality is a key element for stable blast furnace operation, the textural, chemical and physical characteristics of the sinter are very important. Two major factors that affect sinter strength are mineralogy and porosity. This study utilised optical image analysis (OIA) to characterise sinter mineralogy, and the results were compared to those obtained by manual point counting under the optical microscope.Two major mineral characteristics of iron ore sinter are the presence of primary (unreacted) and secondary (derived from melt) hematite. Standard OIA, which is based on RGB thresholding, is not able to distinguish them as they have the same reflectivity; however, primary and secondary hematite in iron ore sinter have a different morphology. The CSIRO-developed OIA software Mineral4/Recognition4 has a unique feature to identify various textures/morphologies of the same mineral. Together with other capabilities of the Mineral4/Recognition4 software, this feature was used to identify different types of hematite and silico-ferrite of calcium and aluminium and other common sinter phases, such as magnetite, larnite, glass and remnant aluminosilicates.A visual comparison of the sinter mineralogy present in optical photomicrographs and that present in the mineral maps created automatically by the Mineral4 module of the CSIRO software confirmed the accuracy of the automated mineral identification.CITATION:Donskoi, E, Hapugoda, S, Lu, L, Poliakov, A, Peterson, M and Haileslassie, A, 2015. Advances in optical image analysis of iron ore sinter, in Proceedings Iron Ore 2015, pp 543–548 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

By Qingwei Wang

This study focused on acidic wastewater from producing sulfuric acid system in zinc smelter containing mercury 21.48mg/L, zinc 55.85mg/L, lead 38.50mg/L, copper 1.68mg/L, cadmium 10.25mg/L, As 29.86mg/L treated by biologics. The factors of pH value, the ratio of biologics to mercury, complexing time, hydrolytic time and temperature were studied in this paper. Experimental results indicated that at pH value 10, quantity of biologics to mercury ratio was 16, complexing time 30min, hydrolyzation time 20min, temperature at 40?, concentration of mercury in treated water was removed to 22.15µg/L. At same time heavy metals such as Zn, Pb, Cu, Cd, As was removed by biologics, the remained concentration in treated water was Zn 0.53mg/L, Pb 0.12mg/L, Cu 0.01mg/L, Cd low to 0.01mg/L and As 0.023mg/L. According to EDAX analysis, mercury content in complexing sediment reached 28.49% easily to recycle as raw material and low concentration of heavy metals in hydrolytic sediment easily to dispose and treat.

Jan 1, 2009

By J. Chipman, J. C. Fulton

Reduction of Si from slag to carbon-saturated iron is a very slow reaction. The rate is nearly independent of stirring but is accelerated markedly by increased temperature. In a slag containing 45 pct SiO2, 38 pct CaO, 17 pct Al2O3, the energy of activation is approximately 130 kcal. AT temperatures of 1450" to 1600°C occurring in the hearth of the iron blast furnace, chemical reactions are normally expected to proceed with great rapidity. Many of the reactions of iron making are indeed quite rapid and attain a condition corresponding roughly to a state of equilibrium before metal and slag are tapped from the furnace. The one conspicuous exception is the reaction by which carbon in the coke or metal reacts with silica in the slag to deliver silicon to the metal. From the practical viewpoint the slowness of this reaction is indeed fortunate, for if it were fast enough to reach equilibrium during the time of contact, it would be difficult to produce pig iron of the desired low silicon content. The one other reaction which under certain circumstances can also be very slow is the transfer of sulfur from blast-furnace metal to slag. In their laboratory exeriments on this reaction, Hatch and Chipman found that at least 5 hr were required for the attainment of equilibrium in a small crucible and that the reaction was not accelerated by increasing the rate of stirring. Subsequently it was shown2 that the presence of a reducible oxide MnO greatly inhibited sulfur removal, and finally that SiO2 itself3 under some conditions has sufficient oxidizing Dower to interfere with sulfur transfer. The kinetics of sulfur removal are now rather well understood,4 and it has been shown3,5 that when the silicon content of the metal is high enough to prevent reduction of SiO,, the desulfurizing process is rapid. Thus the observations of slow removal of sulfur are associated with the coincident slow reduction of SiO2, a process which transfers both silicon and oxygen to the metal phase. It is, of course, the oxygen which interferes with sulfur removal. An explanation of the slowness of the silica reduction reaction has not been obtained. The sulfur transfer process has been shown by the authors5 to be at least partially transport-controlled in the absence of interference by silica. Under the experimental conditions. 99 pct of the total transfer at 1600" was accomplished within an hour. Under similar conditions the reduction of silica may not proceed half-way to equilibrium6 in 8 hr. It is scarcely imaginable that silica reduction could be transport-controlled when the slag contains 44 pct SiO2 and the metal about 1.4 pct C. The experiments reported here show definitely that it is controlled by a slow step or steps in the chemical reaction. One might expect to find such a slow step in the breaking of SiO bonds or in the slow removal of oxygen by carbon at the slag-metal or metal-crucible interface or by bubble formation. EXPERIMENTAL PROCEDURE The experimental procedure used in this investigation was similar to that used in the study of the kinetics of desulfurization.5 A sketch of the graphite crucible and stirrer used in the current work is shown in Fig. 1. The quantity of metal and slag used was adjusted so that the slag-metal interface was located between the top and bottom blades on the stirrer; the top of the slag layer was above the top blade. The stirrer shaft was hollow down to the metal level and served as thermocouple well or optical pyrometer target. Stirring speeds of 0 to 500 rpm were used. A carbon monoxide atmosphere was maintained in the furnace by flushing the gas over the slag surface, except in run No. 220 when it was bubbled

Jan 1, 1960

By Lynne Bowen

"On the da y shift in Number One Mine at Nanaimo on Vancouver Island, on May 3, 1887, William Griffiths, a coal miner with many years experience, fired a shot to advance the face of the crosscut. A flame ran out for a distance of fifteen feet along the roof and then extended up the slope for ten to twelve feet, and although many would have called this a blowing out shot, Griffiths said it was just a feeder. ""A blower"", he said, ""would have caused an explosion."" 1There was no explosion. Griffiths reported gas and the area was cleared for three hours and then work resumed. When the shift changed at 3:00 p.m., the fireman reported all places clear of gas and the ventilation good.2The ventilation in Number One Mine was the best anyone had ever seen. The air in the mine was almost too cold. The current was so strong that it sometimes blew out the flame on the miners' fish oil lamps. It delivered five to six hundred cubic feet of air for every man and animal working in the mine. The law required only one hundred cubic feet 3The miners considered Number One to be relatively safe. The manager was proud of the mine and its up-to-date equipment"". Recently, there had been reports from France and America about the explosiveness of coal dust, but the manager did not think his mine was sufficiently dusty to be dangerous.5 The dust lay thick on the floor of the mine. Sometimes when a shot was fired the dust rose in great clouds. When a miner was shovelling it, the dust would billow up and be carried off by the current of air. The dust was also in the roof and in the stopings, put there intentionally to fill up the spaces and prevent the accumulation of methane gas."

Jan 1, 1992

By G. F. LOUCHLIN

A FEATURE of the International Geological Congress to be held at Pretoria, South Africa, in the summer of 1929,. will be a symposium on the gold resources of the world. In this connection the U. S. Geological Survey and Bureau of Mines have been asked to prepare a report, showing production in the United States, describing the geology of its principal gold- producing districts, and estimating reserves (a) assuming a continuation of presen6economic conditions, and (b) assuming lower costs. This report will not be completed before April 1, but data thus far assembled are here presented with the hope of profiting by discussion and by information from those well informed on any of the important gold-producing districts. Gold production in the United States rose to a maximum exceeding $100,000,000 in 1915 and then rapidly declined to about $52,000,000 in 1920. Since then it has fluctuated, but on the whole slightly declined. Reviews of the eight leading states afford the following tentative conclusions :

Jan 1, 1929

By J. Hallowell, M. Conley, R. Cook

Inflatable friction rock bolts (IFRBs) are often considered temporary ground support due to their susceptibility to weakening from corrosion. The effective service life of an IFRB can be difficult to estimate due to the multitude of factors that affect the rate of corrosion. Some mine sites have experienced several-year-long service lives of IFRBs without issue, while others experience corrosion-related falls of ground only weeks after installation. The most common industry method of assessing bolt performance is by pull testing the fixture. However, due to the IFRB’s ability to mobilize its full tensile capacity with less than 2 feet of anchored length, corrosion along the bolt at depths greater than 2 feet from the collar cannot be identified using this approach. Alternatively, reinflation testing is a simple and non-destructive method of evaluating the condition of an entire bolt length for breaches in the bolt wall. This can be particularly useful when assessing the extent of corrosive conditions within a particular mining area where instabilities have already occurred and the condition of nearby bolts is unknown, or when re-entering historic mine workings which have not been monitored.

Feb 1, 2023

By Eric Toh, Michel David

"Geo-statistics was born from mining problems: grades not being what was expected; more tons of waste than expected; in other words, dilution. Geo-statisticians now claim to have solved many of these problems, but the solutions seem to have a hard time being picked up by practitionners. One of the reasons may be the jargon which geo-statisticians are using. The reason for jargon is that obtaining solutions required a strict formulation in a statistical framework. Practicing mining engineers and geologists are not used to such a framework, and they are not familiar with the jargon. In this paper, the authors re-examine dilution in operating mines and answer two extremely practical problems, using practical language as much as possible. Can dilution be reduced by improving the quality of blast hole samples? How many samples are needed to obtain a good ore-waste discrimination, i.e. to reduce dilution to acceptable levels? Citing real situations, this paper will try to show how fairly simple concepts can be used to answer both problems. It will try to present the solut ions using a minimum of jargon, even if the theoretical solution could be obtained in a few precisely worded paragraphs.The two problems considered are the following: the first one was encountered in the Panguna deposit of Bougainville Copper Limited, P.N .G., but it occurs in every open pit operation where selective mining is required. As the cut-off is lowered, more and more mining occurs in marginal areas. Grade control is mostly based on the grades obtained from blast holes, hence, it is essential that the quality of the blast hole samples be as good as possible in order to reduce dilution, waste being sent to the mill and ore being sent to the waste dump.The second example stemmed from the grade control technique used in many gold open-pit operations in Western Australia: channels are cut on lines 5 m apart and assays are taken every metre. The pattern in the same regardless of rock type, whether one is in the oxidized zone or not. The number of samples to be treated is enormous, hence, the question arises how to find an acceptable sampling grid, and whether changing it can reduce dilution at an acceptable cost.A solution will be given to each problem separately, although in fact it could be seen that the formal framework is the same in both situations."

Jan 1, 1989

By H. C. Stumpf, L. A. Willey, V. L. Fink

Very few studies of the aluminum-sodium system have been reported. Heycock and Neville1 were unable to detect any solubility of sodium in liquid aluminum. Mathewson2 prepared an equilibrium diagram consisting of two horizontal lines at the melting points of the two metals. He was unable to detect solubility at either end of the system. Scheuer3 first developed a method for the chemical determination of sodium in aluminum and then studied the solubility of sodium in liquid aluminum. His method involved heating aluminum with excess sodium under hydrogen until the two liquids were at equilibrium, chilling the melt, and analyzing the aluminum-rich layer. He reported the solubility to be 0.10 pct sodium at 700°C, 0.115 pct at 750°C, and 0.128 pct at 800°C. This paper on the aluminum-sodium system is the twenty-first in a series from the Aluminum Research Laboratories describing equilibrium relations in aluminum alloy systems. Materials The alloys for this investigation were prepared from electrolytically refined aluminum (99.99 pct) and sodium of reagent quality conforming to specifications of the American Chemical Society. Chemical analysis of the aluminum showed the fol- lowing impurities: 0.002 pct Si, 0.002 pct Fe, 0.002 pct Cu, <0.001 pct Mn, 0.002 pct Mg and 0.001 pct Na. The sodium content of all alloys was determined by either chemical or spectro-graphic standard procedures5 No. IIA and No. SI I respectively. Determination of Miscibilitv The boundary of the miscibility gap was determined by a modification of the method of Scheuer3 and of that used by Henry and Cordiano. As finally developed the method was as follows: A fused alumina crucible 25 mm id with 75 mm depth and 1.5 mm wall thickness was placed on the bottom of a container made from a 4-in. length of 1 1/2-in. iron pipe and pipe caps. ?&apos;he space around the sides of the crucible was filled with 40 mesh crystalline alun-dum. A charge of 60 g of aluminum and 10 to 15 g of sodium was placed in the crucible, the sodium first being trimmed and rinsed with iso-pentane to remove kerosene. The container was tightly capped, filled with argon (99.6 pct pure), and heated to the desired temperature (between 670 and 800°C). A positive pressure of argon (about 8 mm of mercury) was maintained through a small pipe to the upper cap. After soaking for from 2 to 5 hr, the container was quenched by directing a stream of water onto the base, so that solidification of the aluminum-rich layer was rapid and essentially unidirectional. Analyses were made on drillings taken from the bottom of the ingots after machining off the surface layers. Pre-

Jan 1, 1949

By John J. Dwyer

As one begins to contemplate the organization of a talk with the title "Iron Ore - The Permanent State of Evolution," the main divisions appear to be the past, the present, and the future. But distaste for repetition of the present production achievements, which scarcely need additional description on this occasion, soon leads one to the conclusion that a more appropriate division of the subject is simply the past and the future . . . with emphasis on the future. We are working in the present, but it can only be understood by what has taken place in the years and decades gone by, and/or under what terms we will have to face the future. Evolution is more often thought of as a steady change - but steady change is really the product of revolutions, sometimes making headlines but often unnoticed at the time. With hindsight, we recognize the great importance of the emergence of the Bessemer converter and the open hearth and the vast influence that both have had on steel making and the changing ore demand. In this country we also see schooners opening the era for lake transportation of iron ore from Michigan to Take Erie. Later, steamships haul the ore from the Mesabi Range to the steel centers of the country. In order to take full advantage of the inland seaway, there had to be revolutions in ship design, loading and unloading facilities and channel and lock development. Full scale utilization of the resources of the Mesabi Range, itself, required parallel upheavals such as the development of new production equipment, and substitution of modern forms of energy for horsepower, and, yes, manpower, as well.

Jan 1, 1968