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By Charles G. Maier

"INTRODUCTION The use of thermodynamic calculations to answer practical problems in metallurgy, especially those that are relatively difficult to test experimentally, is rapidly becoming a conventional method of procedure. Nevertheless, these methods cannot he applied effectively until the fundamental thermodynamic constants used for them have become available either by direct experiment or by recalculation of data already existing in more or less undigested form.In an attempt to collect such data far the compounds of lead and the antimony group the incentive has been to make a preliminary survey that will disclose the gaps or questionable phases of available information and to collect data that may be used to answer certain rather practical problems presented by operating metallurgists.This report comprises three sections; the first gives the basis of selection of the constants which may now be given or estimated, and the later sections discuss the application of the figures obtained to two of the more interesting problems.The reader will discover that many of the quantities involve critical study of a number of more or less discordant observations. Although certain data could be rejected on the basis of thermodynamic comparisons, it must not be overlooked that many of the figures are the results of rational estimates rather than exact experiment. The use of thermodynamic methods to control and partly authenticate such estimates is one of its most valuable characteristics for the practising metallurgist or chemist, who cannot check by direct experiment each property of the materials which he converts to usable form.Nevertheless, while it is definitely possible in certain instances that the final figure selected for the heat or free energy of a process or substance may be in error by as much as a few hundred to a few thousand calories per formula weight, questionable figures are usually revealed by the thermodynamic calculations made. When quantities of such degree of accuracy are, and they often must be, used for estimating an unknown reaction, the result often gives no more than the order of magnitude of the desired quantities. Yet even in these extreme cases the answer is far more reliable and valuable than an unordered guess or a conclusion based on incorrect data."

Dec 1, 1934

By Walter R. Storms

"The New Mexico potash exploration described herein was begun in February- 1944 by the Federal Bureau of Mines with the object of determining the amount of potash ore in the United States Potash Reserve in Eddy County, N. Mex., by exploring the deposit from the surface with vertical drill holes. The work was completed in October of the same year after 16 holes had been drilled on the reserve and in the immediate vicinity.The Potash Reserve is located in secs. 26, 27, 34, and 35, T. 20 S., R. 29 E., Eddy County, N. Mex., about 14 miles northeast of the city of Carlsbad. Drilling was done in secs. 26 and 35 of the reserve, in sec. 14 to the north, and secs. 25 and 36 to the east. All work was done on Government-owned, unleased land.The reserve may be reached from Carlsbad by traveling northeast by automobile for about 14 miles on paved U. S. Highway 62. This highway passes through the southern part of the reserve. Other parts may be reached by a State highway which passes along the eastern side of the area, then by poor dirt roads to various drilling sites. A branch of the Atchison, Topeka and Santa Fe Railroad also passes across the southern part of the reserve en route from Carlsbad to the plant and mine of the Potash Co. of America, a few miles northeast of the Project and 22 miles from Carlsbad.Work on the project was greatly facilitated by the help of Ralph H. King, associate geologist, R. K. Bailey, geochemist, H. C. Spicer, geophysicist, Hope Bunker, junior geologist, and R. H Allport, district mining supervisor, all of the Federal Geological Survey; and Charles H. Johnson, who at that time was district engineer for New Mexico for the Mining Branch of the Bureau of Mines."

Apr 1, 1946

By L. H. McGuire, L. B. Berger

"During the period October 31 to November 4, 1944, tests were conducted by the Bureau of Mines in the Cascade Tunnel of the Great Northern Railway Cc. to determine the effects upon the tunnel atmosphere of the use of a Diesel freight locomotive through the tunnel. These tests were conducted at the request of the company to ascertain whether undesirable conditions might be caused by the discharge of Diesel engine exhaust gas in the tunnel, which was constructed for an electric haulage system and is not provided with equipment for mechanical ventilation. The use of Diesel locomotives on the division that includes the Cascade Tunnel was under consideration as a means of augmenting haulage facilities for the abnormal freight tonnage that must be handled during wartime.As the Bureau of Mines has received requests for information on the extent of contamination of the air of railroad tunnels by Diesel locomotives, this report was prepared to make available for general information the data obtained during the tests in the Cascade Tunnel. As the test locomotive is of a type used on other railroads, and as the bore and grade of the tunnel and the tonnages moved are comparable to such factors on other roads, the data should be sufficiently representative to be of general interest. The exceptional length of the Cascade Tunnel made it an excellent location for obtaining information on the composition of the tunnel air and on the movement of air induced in the tunnel by the passage of trains.The writers are indebted to C. O. Jenks, vice president of the Great Northern Railway Co., who arranged for the tests, and to the following personnel of the railroad who cooperated in the conduct of the tests: I. E. Clary, superintendent, Spokane Division, J. L. Robson, superintendent of motive power, oil and gas equipment, J. F. N. Gaynor, superintendent of electric operations, William Lowney, master mechanic, Spokane Division, and D. F. Marston and Verne Nelson, assistant superintendents of electric operations.The Cascade Tunnel is on the Spokane Division of the Great Northern Railway and pierces the Cascade Mountains some 49 miles west of Wenatchee, Wash. The west end of this division is electrified from Appleyard Station, 2 miles east of Wenatchee, to Skykomish, Wash., a distance of 73 miles. Until the described tests, only electric haulage had been employed through the tunnel since it was first opened to traffic in January 1929.The Cascade is the longest railroad tunnel in the United States - 7.79 miles. The tunnel is single-track, perfectly straight, concrete-lined, and is on an essentially uniform grade of 1.56 percent, ascending from west to east. The cross-sectional area of the tunnel is 314 square feet, the dimensions being shown in figure 1. Elevation of the east portal is 2,880 feet and of the west portal, 2,240 feet."

Jun 1, 1946

One of the chief drawbacks to Bendigo mining always has been the difficulty of draining the mines in an equitable manner, so that each company or mine owner bears the properly share. The trouble seems to have been apparent on this field since the diggings were discovered. One mine would push on more expeditiously than its neighbour in looking for the precious metal and sink its shaft, whilst those in the vicinity would simply potter about, and discontinue sinking until gold was struck in the deeper ground, and: then they would go ahead, start sinking, while the pioneer would have to pump all or most of their water. This sort of thing continued here for years until the leases were worked down to the very deep ground and the water became such a b.urden on those companies that had deep shafts, that first one stopped hauiing and then others; with the result that over 1000 feet of water was allowed to accumulate on the New Chum line. lt appeared that if something were not done the whole of the main line of reef in the Bendigo district would be at a standstill, when a scheme was formulated and handed to Mr. Lansell, who, with other gentlemen, took the matter up, and at the present time all the mines on that particular portion of reef are in full work and the water all bailed out.lt is this scheme that I am requested to place before you and also to give you the details of the work done since its adoption.

Jan 1, 1902

By Cadwallader Evans

MY endeavor in this paper will be to tell something of the accomplishments and current problems of the anthracite producer and to suggest some of the avenues for technical development that seem to me to be the most necessary and likely to bear fruitful results. The several subjects, that I find of most interest will be taken up individually with brief reference to the future trend. Underground Mining and Loading ?In this field great progress has been made in the last twenty years, beginning with the introduction of the scraper system. The electrical hoists for the operation of scrapers have been improved, and give little trouble, but there is an open field for the improvement of the auxiliary apparatus, particularly in the application of roller-hearing sheaves to the guiding of scraper ropes. Also notable in the last six or seven years is the introduction of shaker chutes, so that far more anthracite is now loaded over them than by scrapers. This trend should continue to be manifest, particularly if those who are looking ahead and who are charged with the development of this specialty continue constantly to study and improve the mechanical design of their shakers.

Jan 1, 1938

By Andrea Bernardes

Printed circuit boards (PCBs) of varying compositions have been converted by incineration and following melting into an environmental agreeable slag and a copper-nickel-tin alloy, containing the precious metals. The environmental compatibility of the slag was determined according to the German standard DIN 38414 - Part 4; pollutants' concentration in the lixiviates is smaller than the thresholds of drinking water (CE-Standards). Each experiment the charge were 500g scrap from PCBs and ca. 100g slag forming material. The products were a kind of mullite slag, poor on iron, which was recirculated, and an alloy containing up to 0.3% Gold. With a burner to the retort coal gas, the waste gas was burned without soot, and the produced mixed zinc-lead oxide containing silver was collected in the flue dust.

Jan 1, 1997

By J. van der Merwe, D. Tharandt

The corrosion behaviour of 304L stainless steel laser-cladded with various amounts of ruthenium (Ru) was evaluated in solutions of sulphuric acid and sulphuric acid plus sodium chloride at 25°C and 45°C by open-circuit potential and cyclic potentiodynamic polarization tests. In general, the addition of Ru to the stainless steel increased its corrosion resistance in 1 M H2SO4, as well as in 1 M H2SO4 plus 1 wt% NaCl. This was observed for a number of parameters such as corrosion rate, corrosion potential, open-circuit potential, and current density. However, increasing the amount of Ru added beyond a certain level did not result in further improvement in corrosion protection. For each environment there is an optimal Ru concentration for the best corrosion protection. For example, in 1 M H2SO4 at 25°C, 2.44 wt% Ru shows the least active surface in terms of corrosion. Further research into ruthenium coatings on stainless steels is recommended.

Jan 1, 2015

By T. P. Mucho

Strata behavior and support performance were evaluated in a longwall tailgate test area at the Eagle Nest Mine near Van, WV. The mine operates in the Eagle coalbed which ranges in height from 4 to 6 ft (1.2 to 1.8 m) on the mine property. Rugged topography results in rapid changes in overburden which ranges from 200 to 1200 ft (61 to 366 m). The immediate roof at this mine transitions from sandstone to shale. In some areas the sandstone appears to be massive while in many locations it is highly laminated, fossilized, and interspersed with coal streaks. Horizontal stress levels appear to be sufficient to create instabilities in the roof in some locations, especially where the roof is thinly laminated. Traditionally, hardwood cribs have been used at the mine to provide secondary support for longwall tailgates; Strata Products' Hercules cribs are installed on 8 ft (2.4 m) centers staggered left and right in the tailgate entry. In a 250 ft (75 m) tailgate test area, however, cable bolts were installed in lieu of cribs. The bolts were 12 ft (3.6 m) long tensionable cables anchored with 5 ft (1.5 m) of resin. Four cables were installed across the entry and rows were spaced on 6 ft (1.8 m) centers. Primary roof support was maintained in the vicinity of the test area using 42 in (1.1 m), grade 60, No. 6 (19 mm diameter) resin bolts installed on 4 ft (1.2 m) centers through T-2 channels. The cable test area was located under a stream valley in an area of relatively shallow overburden. Strata instabilities consistently had been associated with the stream valley in both the Eagle and the overlying Powellton coalbeds. In an effort to further expose the cables to the most difficult stress conditions available at the time of the test, the specific site chosen was directly beneath a barrier pillar remaining in the upper Powellton mine (approximately 150 ft (45 m) of interburden separates the two coalbeds).

Jan 1, 1996

By Dennis R. Dolinar, John L. Ellenberger, Gabriel S. Esterhuizen, Anthony T. Iannacchione

A survey of pillar conditions was carried out at 21 operating limestone mines that use the room-and-pillar method. The surveyed pillars were all located in rock that was classified as “Good” to “Very Good” using the RMR rock mass classification system. The average pillar width was 14.5m and the width to height ratio varied between 0.53 and 3.52. Pillar conditions were rated using a visual rating system that accounts for the effect of geological structures as well as stress related fracturing. Nine cases of pillar instability were recorded. Pillar instabilities were typically limited to individual pillars or small groups of pillars within an otherwise stable layout. Local geological structures such as inclined discontinuities or weak bedding planes contributed to instability in seven of nine cases that were observed. Elevated pillar stresses contributed to instability in the two remaining cases. All the unstable pillars observed had width-to-height ratios of less than 1.5. It is concluded that pillar instability is most likely to be caused by unfavorable geological structures in pillars with width to height ratios of less than 1.5. Stress related instability such as rib spalling becomes more prevalent when the average pillar stress approaches approximately 20 MPa.

By Wern-Ping Nick Chen

Durability risks for underground facilities can be underestimated during design phase as compared to well-known geological risks. However, unclear durability requirements can result in construction schedule delay and change orders. Unlike bridges or buildings, a unified underground durability code or standard does not exist. This paper will provide a durability approach for underground facilities, including degradation modes, mitigations, and compliance measures during construction. In addition, this paper will address how to clearly specify durability requirements in a contract document to reduce disputes during construction and discuss key durability tests that are time consuming and will impact construction schedule if not timely planned.

Jun 23, 2024

By D. Brand

"The risk management objective of the AMIRA Code of Practice for Metal Accounting has seldom been systematically investigated in the literature up to the authors’ knowledge. Risk is defined in the Australian Standard on Risk Management (AS/NZS 4360:2004), as “the chance of something happening that will have an impact on objectives”. The simplified economic analysis performed in this paper has lead to define “something happening” as measurement non-representativity and “impact on objectives” as NSR loss from the optimal NSR. This has pinpointed a missing link being the establishment of a relation between NSR in flotation plants and measurement representativity. The proposed framework, referred to as the total cost of representativity, answers the previously unanswered question of how much accuracy/reproducibility is fit for purpose in a given flotation plant. Authors readily acknowledge that it is way easier to assess the cost of accuracy/reproducibility than the cost of inaccuracy/irreproducibility. However, this exercise not only answers previously unanswered questions but also establishes an essential link with operational excellence. This paper exposes why such operational excellence could neither be attained nor proved without a state of the art metal accounting system. This should impact not only how metal accounting systems are audited/designed/refactored, but also how flotation plants are designed.IntroductionBecause representative metal accounting is a vital source of information for the financial accounts of a mining/metallurgical company, the need for a code on metal accounting was recognized in 2001 and a team was formed in 2003, under the sponsorship of several major mining houses, to produce such a code. Table 1 exposes the two main objectives of this AMIRA Code of Practice for Metal Accounting."

Jan 1, 2017

By Reinhardt Thiessen

Views and conceptions of the origin, composition, and general nature of coal differ so widely that to determine the real extent of actual knowledge is no easy matter. The chemist finds difficulty in attacking the chemical problems and the engineer lacks a broad chemical basis for the study of his problems. Thus the economical and efficient utilization of coal in the industries suffers from a lack of knowledge of the composition of coal itself. Most scientists interested in this matter, as has been brought out in discussions during the recent investigation of coal problems, appear to have only a vague idea of the components of coal and their origin. Some are even in doubt as to whether substances derived from the wood and resins of plants form a part of coal; and few seem to have more than the dimmest idea of what are the components of coal, the condition in which these now exist, and the transformations that they haye undergone. Although during the paiilt century or more many co.rrect observations have been made and recorded, these were falsely or inadequately interpreted, so that many of the interpretations rendered are contradictory. As a result there has been great confusion. Many of the observations made in the past, although accurate and correct, could not be explained, because the principles involved in a correct solution of the problem were not then known. For example, Parkinson, Hutton, Petzhold, Guembel, and others made valuable discoveries regarding the processes by which plant substances are transformed into coal, but as the principles involved are those of colloidal chemistry, a science not known to the scientific world of that time, the observed facts could not be interpreted and the solution of the whole problem was blocked. The explanation of the true meaning of the alternating bright and dull bands so familiar in all ordinary bituminous coals afforCls another example. This banding has been observed and recorded from the earliest times, and has led to the forming of many false theories. Some gAologists, for example, have endeavored to explain the bands as due to alternating climatic conditions during the coal-forming period. Had the geologists known the true nature of coal they could not have made this mistake nor many others made in the past. To fill in a few of these gaps and to clarify some of the existing confusion is the aim of this paper. The immediate cause that led the author to investigate further the composition of coal and the origin and structure of its constituents was that when an attempt was made to discover the meaning of the coking and noncoking properties of coal a more exact knowledge of the nature of coal was required than was then available.

Jan 1, 1920

By Charles E. McCarthy

Part 1 of this report is concerned with the electric-furnace production of calcium aluminate and iron alloy from a variety of siliceous ferruginous ores of aluminum. The furnacing research was directed, first, to determine the extent of the field in the CaO-Al2O3-SiO2 system, in which calcium aluminates with the optimum leaching characteristics are obtained; second, to produce suitable quantities of these aluminate mixtures for pilot-plant leaching tests; third, to investigate the physical chemistry of the reduction reactions during furnacing; and, fourth, to study the economics of the process. The third and fourth steps were not completed. Part 2 comprises another phase of the work and will appear as a second Report of Investigations. It deals with the extraction of metallurgical-grade alumina from the calcium aluminate slag produced in the electric furnace and includes the following steps: leaching, desilicating, precipitating, dewatering, filtering, and washing of the aluminum hydrate. No claim is made that the data are complete, nor has the chronological order of obtaining data been followed. Rather, it is emphasized that additional work is necessary if the full possibilities of the process are to be evaluated. and made available.

Jan 1, 1949

By L. Mabiza

Keywords: Pyrometallurgy, furnace, sulphur, SO2, off-gas, Zimplats Zimplats embarked on an off-gas sampling campaign in December 2009 to characterise the concentration of SO2 in the off-gas from their 13 MW furnace. In excess of forty isokinetic samples (US EPA Method 8) were taken over seven days. The aim of the sampling campaign was to increase the understanding of factors that influence sulphur release from sulphide concentrates in an electric furnace. Specifically, detailed monitoring was conducted over various operating conditions to evaluate: ? The relationship between SO2 and feed rate (i.e. furnace power) ? The relationship between SO2 and ingress (i.e. furnace freeboard pressure) ? The impact of concentrate bed thickness on SO2 (i.e. blacktop) ? The effect of events such as tapping and converter slag return on SO2 ? The desulphurization reactions that take place in the blacktop A model was developed to predict the SO2 content in furnace off-gas for a particular set of operating conditions. In order to evaluate the impact of random, uncorrelated changes in concentrate composition, ingress and power (feed rate) on the SO2 content of the off-gas, a Monte Carlo approach was adopted whereby concentrate composition, power, and ingress were allowed to vary randomly around selected set-points. The model is primarily based on the release of ?labile? sulphur from the partial thermal dissociation of sulphide minerals with an over-stoichiometric sulphur content (e.g., Fe7S8 decomposition to FeS and S). The main findings of this work suggest that labile sulphur does not fully account for the measured SO2 content of the off-gas, and that limited oxidation of sulphides during melting also contributes to the SO2 content. Details of sulphide oxidation reactions in the furnace remain unclear, and it is currently not possible to provide a theoretical framework for the relationship between the extent of oxidation and variables such as temperature, gas composition, particle size, and residence time in the furnace black top. Through the use of the Monte Carlo model it was, however, possible to constrain the extent of sulphide oxidation to about 1%. This work has shown in practice what was expected in concept, that SO2 in off-gas is proportional to furnace power and ingress rate. The model that was developed consistently slightly under-predicts the quantity of SO2 generated. It also highlights that more work is required to better understand the analytical methods used to analyse for SO2 in off-gas.

Jan 1, 2011

By R. E. Ray, J. W. Stevenson, J. A. Berry

Conventional concrete block plastered with a cementitious coating is the most common material used in the construction of permanent stoppings to direct airflow in underground mines in the US. All mines experience various degrees of strata convergence depending on depth of overburden, geological conditions, and type of roof support employed. Strata convergence will cause cracks and joint openings in masonry stoppings, resulting insignificant air leakage losses. Where strata convergence is severe, complete structural failure of the stopping can ultimately occur. Reconstruction of damaged or destroyed stoppings adds expensive overhead to mining operations, and even greater expenses are incurred from the additional fan horsepower required to overcome leakage losses. Ideally, a stopping should maintain high resistance to airflow while yielding to strata convergence. By properly incorporating a polyisocyanurate rigid foam material within the masonry block structure, stopping service life can be increased in mines experiencing strata convergence problems such as floor heave, roof loading, and lateral rib movement.

Jan 1, 1986

By C. M. Harrer

The location, features, qualities, and extent of iron occurrences and the associated resources--limestone and dolomite, bentonite, coal-lignite, petroleum and natural gas, power, and water--are evaluated. The results of pioneer beneficiation tests on South Dakota taconite are discussed. In addition, pertinent features of taconite development, cost factors, the Western and Central United States trade area and markets are reported. The State's production of iron ores until 1963 has been small and for purposes other than the manufacture of iron. This investigation disclosed the probability of large resources of taconite in Pennington and Lawrence Counties. The average iron content is 29 percent, and a 65-percent or better iron concentrate might be produced. Ample resources are available for mining and processing taconite. An investment of at least $26 million would be required for a plant capable of turning out 1 million tons of agglomerated iron concentrate per year. The feasibility of such an enterprise is indicated, but intensive investigation and research must be completed to justify such an investment,

Jan 1, 1966

By Michel Brissette

"To recover more valuable ore and considering the complexity of minerals, the needs to grind finer is increasing. To reach the desired fine grinding targets, using the existing grinding media size, the energy consumption is increasing exponentially. As already proved at this conference, small grinding media represents the best potential to improve grinding efficiency. How do the small grinding media perform in industrial grinding mills?The use of small grinding media in regrind mills proved that finer grinding can be achieved at lower energy consumption. In ball mills, smaller grinding media versus 25 mm media generate a power saving from 10% to 44%. In vertical stirred mills, the power saving increases from 20% to 60%. More potential savings have been identified.For the same operating conditions, how a vertical stirred mill is performing compared to ball milling? In regrind application, a vertical stirred mill with small media (5-12 mm Millpebs) will require at least 62% less energy than a ball mill charged with 25 mm grinding media.INTRODUCTIONIn some mining applications, fine grinding is needed to liberate further the valuable ore. The purpose of regrind mills is to achieve the last stage of grinding before final flotation stage: the cleaners. In part because of high prices of small grinding media as well as availability, the most widely used size of grinding media is 25 mm.When calculating the recommended ball top size from Bond’s formula (Bond, 1961), the size for regrind applications is much lower than 25 mm. Today, the barrier for small grinding media is broken. They are best suited in regrind application. The size of grinding media is one of the most important variables affecting the mill efficiency (McIvor, R., 1997; Nesset, J.E., Radziszewski, P., Hardie, C. & Leroux, D.P., 2006).Conventional ball milling was discarded for fine grinding because of its very poor energy efficiency (Zheng, J., Harris, C.C. & Somasundaran, P., 1994; Yan, D.S., Freeman, M. & Dunne, R., 1995). Therefore, new technologies as stirred mills were developed to grind finer more effectively. Three fundamental questions should be asked. What is the best fine grinding technology? What should be the best media size to use? Is it possible to grind finer by using less energy with the existing regrind equipment?"

Jan 1, 2009

By Lawrence Pelham

Introduction Fluorspar demand or consumption has been taken for granted in recent years, except for those that produce, market, or consume the commodity. Thinking used to be that as the steel and aluminum industry goes, so goes fluorspar. This is much less true today. The industry is becoming more dependent on chemical industries that feed primarily off hydrofluoric acid (HF) production. The US Bureau of Mines (USBM) published an assessment of world fluorspar reserves and resources. It created a foundation of valuable data that can be used to monitor fluorspar availability. Celebrations were held this year commemorating the 100th anniversary of the isolation of fluorine. The US and other countries continue to struggle with the controversy of ozone depletion and the effect of fluorocarbons on the atmosphere. Goals for the government's strategic stockpile are being reevaluated. The market for fluorine chemicals is changing and diversifying, making it a challenge to assess and forecast fluorspar/fluorine demand. US production and distribution Fluorspar Fluorspar shipments from US mines reached an all-time high near the end of World War II in 1944 at 375 kt (413,800 st). Eighty-two mines produced more than 453 t (500 st) each, and 10 mines produced more than 9 kt (10,000 st.) each. Illinois supplied 43% of total shipments. Consumption in 1944 was 372 kt (410,200 st), of which 56% was used by steel producers and 32% by HF producers. This was the last year that US production exceeded US consumption. The post-World War II record for fluorspar shipments occurred in 1951 at 315 kt (347,000 st) 58 mines produced more than 453 t (500 st) and eight mines produced more than 9 kt (10,000 st).

Jan 11, 1986

By T. T. Tomimatsu

This? Bureau of Mines paper discusses the corporate structure dynamics, methods of financing, and financial assessment approaches that should be utilized to evaluate the economic health of the coal industry. It suggests rates of return and other measures of corporate economies necessary to attract capital required to finance emerging coal demand, plus impact of coal sales to total revenues. The study illuminates the activities of 30 selected coal-producing companies, including their subsidiaries or affiliates, that were responsible for approximately 60 percent of the total U.S. production in 1974. The selected firms are not all primary coal producers; many of the larger producer's are classified by the financial institutions as petroleum, metal, and steel producers, public utilities, or chemical industries. The bituminous coal industry is highly competitive as to price, service, and quality of pronuct and is subject to competition from other noncoal energy sources. The top 15 bituminous coal-producing companies mined about 46.6 per-cent of the Nation's total production in 1974, and the balance of approximately 3,885 companies were responsible for the rest of the production. Thus, the U.S. bituminous coal industry is not an oligopoly. Today, about 65 percent of the Nation's total coal production is consumed by electric power utilities and 26 percent by steel, space heating, and other industries; the balance of 9 percent is exported.

Jan 1, 1976

By W. B. Boggs

The Noranda smelter was built by Noranda Mines, Limited, from the designs of A. E. Wheeler, consulting metallurgist, New York, for the reduction of ores and concentrates from the Horne mine, and to treat, on a custom basis, ores and concentrates from other mines in the Rouyn district. The smelter is situated at Noranda in western Quebec, about 330 miles directly north of Toronto. Ground was first broken for construction in November, 1925, before the completion of the branch railway into the district, which was not finished until October, 1926. After many difficulties and delays due to the Jack 'of transportation and the severe weather during the winter months, the smelter was started in December, 1927, and the first blister copper was poured on December 17th. The smelter was built throughout in two duplicate units, each with a rated capacity of 500 tons of ore per day. Operations were started with only one unit, and after several months it was found that one unit alone would smelt 1,000 tons of ore per day. At this time only one shaft, No. 3, existed, and its hoisting capacity and the capacity of the original crushing plant were insufficient to supply ore for two units at the increased tonnage. Accordingly No. 4 shaft was sunk and a new crushing plant provided to handle the ore hoisted there. This work was completed in the latter part of November, 1929, and since that time both smelter units have been operating.

Jan 1, 1930