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By V. K. Karra

A mathematical model for the performance of a vibrating screen is developed in terms of an oversize partition curve. This is a normalized curve for any assemblage of particles presented for screening when their size is normalized with respect to the 50 per cent passing size (d50 ) under the prevailing screening conditions. Based on the throughfall aperture (hT) of the wire mesh on the deck, the effects of oversize (% >hT), half-size under (% <0.5 hT), deck location and feed rate were evaluated using the existing screen sizing information. All these variables and also the near-mesh (1.25 hTx 0.75 hT) content (%) of the feed affect the d50. Performance data from different decks of a double-deck screen with various feed rates, size distributions and screen apertures were used to evaluate the near-mesh factor. Effects of wet screening and material bulk density are also incorporated in the model. This generalized model is useful for the simulation of a multiple-deck screen. It furnishes the sieve analysis information of all the deck output streams and therefore has a potential use in the over-all circuit evaluation when used in conjunction with relevant crushing and grinding performance models. An example illustrating the prediction capabilities of the model is given.

Jan 1, 1979

By E. L. Garton

The United States Bureau of Mines has recently collected samples of crude petroleum from wells in the newly developed Joiner. Kilgore, and Longview fields of Rusk and Gregg Counties in the East Texas district, and from the Van field of the East Central district of Texas.

Jan 1, 1931

By Steve Kral

One of the world’s most successful mining shovel manufacturers believes its name is not very well known in North America, particularly in the United States. And there are plans under way to improve on that. Never mind that it is one of the largest mining shovel manufacturers in the world. Or that the mining division has 800 machines in operation worldwide, 100 of them in the United States. And the company currently holds 30% of market share worldwide and 10% in the United States. Forget about the fact that it has already sold more machines world-wide this year than it did during 1995. None of that is good enough. O&K Mining GmbH Dortmund is the mining division of Germany’s giant O&K Orenstein and Koppel. It is one of three divisions of O&K Orenstein and Koppel AG, a US$1.2-billion company. The other two divisions are construction equipment and escalators. The mining division, though, is one of the company’s most successful.

Jan 1, 1996

By João Batista Ferreira Neto

A escória de aciaria pode ser uma alternativa como mistura para o cimento portland, substituindo parcialmente a escória de alto-forno. Um processo pirometalúrgico tem sido desenvolvido para promover a modificação química da escória de aciaria LD. Os experimentos foram realizados através da refusão de 300 kg de escória de aciaria LD, seguido de sua modificação no estado líquido. Essas escórias foram resfriadas com o uso de esferas metálicas. Análises de fluorescência de raios–X (FRX) e microscopia eletrônica de varredura (MEV) das escórias revelaram uma relação entre composição química, condição de resfriamento e a formação de fases cristalina ou vítrea. As amostras de cimento foram obtidas através da mistura de 25% de escória modificada e 75% de cimento Portland, resultando em mais de 280 J/g de calor de hidratação acumulado em 72h, expansão abaixo de 0,1% em ensaio em autoclave e resistências à compressão acima de 42 MPa com 28 dias. O processo mostra potencial para ser aplicado no tratamento de escória de aciaria LD.

Aug 17, 2017

By Foluke Salgado de Assis

Este trabalho visa apresentar alguns resultados obtidos através de ensaios balísticos realizados em sistemas de blindagem multicamada. Seguindo as especificações da norma americana NIJ 0101.06, foram realizados ensaios utilizando munição de calibre 7,62x 51 mm, que possui velocidade acima de 800 m/s. A blindagem multicamada utilizada possui como camada frontal uma pastilha hexagonal cerâmica (alumina dopada com 4% de nióbia), segunda camada de compósito de matriz poliéster reforçada com tecido de juta e como terceira camada uma placa de alumino 5052 H34. A utilização de compósitos poliméricos reforçados com fibras naturais em substituição às fibras de aramida é de grande interesse, pois além de desempenharem a mesma função no sistema de blindagem, possuem um valor de mercado bem menor. Com os resultados obtidos do ensaio balístico, os valores foram associados com o seu comportamento dinâmico utilizando o método do casamento de impedâncias.

Aug 17, 2017

By M Nehring, M Kizil, P Knights, A Mousavi, F Sotoudeh

A cut-off grade is used to distinguish an economically viable block of ore from non-economical ones in mining operations. In order to obtain a high Net Present Value (NPV), a strategic and precise determination of cut-off grades (mill feed/marginal, marginal/waste) is required. In recent years, the significant depletion of feed grades and increasing mining operating costs have affected the viability of new and existing mining operations. Hence, the introduction of a pre-concentration system to process ore before sending it the main processing plant has gained increasing attention for its potential improvements in resource exploitation processes, mineral processing, energy consumption and environmental impacts. The primary potential advantage of implementing a pre-concentration system is the improvement of resource utilisation and increased reserves as a result of a reduction in cut-off grade. The objective of this paper is to present a new cut-off grade method, which can be used in the presence of a pre-concentration system for underground metalliferous mining operations. The results of this study illustrate the potential economic benefits of using a pre-concertation system prior to a mineral processing operation as a result of the reduced cut-off grade and increased NPV.

Nov 10, 2020

By S. W. Parr, T. E. Layng

There is presented below the results of experiments made on six samples of Utah coals furnished through the U. S. Bureau of Mines , under an agreement dated June 16th , 1920. The details of all values for all runs are tabulated and presented herewith . Clear The samples, consisting of approximately 150 pounds each, were received at the University of Illinois , on June 15 , 1920 , and are listed as follows : Creek , Wattis , Castle Gate , King , Standard , and Aberdeen . The proximate and ultimate, analyses of the various ccals from that region , representing all the types included in these samples , have already been made by the Bureau of Mines , and are available . * Similar determinations were also made on the particular samples forwarded for purposes of these experiments . The results obtained on the coal as received did . not very in any material feature from the values reported by the Bureau, they are shown in Table 1. It is of importance to note only that the oxygen content is uniformly high , namely approximately 9.5 to 11.5 per cent of the as- received coal , while the oxygen - hydrogen ratio is also high, being on the average approximately 2 ( oxygen) to 1 ( hydrogen) .

Sep 1, 1921

By P. Kim, M. Tangstad, T. Larssen, R. Kawamoto

The reduction rates of MnO and SiO2 were investigated for various charge compositions. The empirical activation energies of MnO and SiO2 reduction from SiMn slags were calculated between 1500 and 1650 °C under CO atmospheric pressure. The amounts of metal produced at different temperatures were compared, and the Arrhenius plots were described for each slag A, slag B and synthetic slag, which shows that the raw material have significant influence on the reduction rate. The rate of Mn-produced in slag A was faster than slag B despite of the relatively lower driving force. The lower rate constant in slag B implies that other kinetic factors, such as viscosity, affects the reduction rate. The estimated activation energies of MnO were 920, 304 and 975 kJ/mol MnO for slag A, slag B and synthetic slag, respectively. The rates of Si-produced were similar and slow for all slag types. However, the different activation energies of SiO2 between industrial ores and synthetic materials implied that the trace elements can have catalytic effects toward SiO2 reduction in SiMn slags. The estimated activation energies of SiO2 reduction were around 800 kJ/mol for slag A and slag B.

Mar 1, 2017

By De-qing Zhu, Feng Zhang, Jian Pan

Sintering performance of a series of sinter blends with high ratio of limonite was investigated. The metallurgical performance and mineralogy of product sinter were also examined. The results show that good sintering performance could be obtained when the ratio of limonite was below 43% in the blends of iron ore fines, while good microstructure and metallurgical properties were achieved under optimum sintering conditions. In contrast, sintering performance will rapidly deteriorate when the ratio of limonite is elevated up to 63% in the blends.

Mar 1, 2017

By Jiongtian Liu, Guihong Han, Shuzhen Yang, Yanfang Huang

Binder, as a necessary functional material of electrode in lithium ion battery, has a vital influence on the electrochemical performance. The humics, coupled with polyacrylamide (PAM) and sodium carboxymethyl cellulose (CMC), were used as a water-soluble binder for LiFePO4 (LFP) cathode using in lithium ion battery. The techniques of cyclic voltammetry (CV) and galvanostatic charge-discharge were conducted to evaluate the performance of humics binder in LFP electrode. The reversible specific capacity of LFP (HA + PAM + CMC) electrode can reach to 142 mAh g−1 after 50th cycling (100 mAh g−1) on conditions of raw materials with 1 wt% PAM addition, humics and CMC with a mass ratio of 3:1 in binder, and the raw materials mixed in an interval order. Humics has the potential to be used as a novel, eco-friendly and high-efficient binder for lithium-ion battery.

Mar 1, 2018

The Bureau of Mines in the Department of the Interior is increasingly concerned with the safety and health of personnel engaged in coal-mining operations, particularly below ground. The Bureau is also concerned with improving the utilization of domestic mineral resources, and reducing the environmental impact of mineral operations. One obvious way to improve safety and minimize health hazards would be to eliminate coal mining entirely by bringing the energy value of the coal to the surface in some other way, such as by partially burning it underground to form combustible gases and bringing these gases to the surface, where their energy content can be put to use. This technique might also improve the utilization of the coal resource. ? Accordingly, the Bureau of Mines, interested in re-evaluating underground coal gasification technology in terms of current and expected conditions in the United States, requested Arthur D. Little, Inc., (ADL) to appraise the technology to determine: ? Whether the current (1971) state of the technology offers a technically feasible method for gasifying coal underground that is applicable to the reserves of bituminous and lower rank coals in the United States. ? Whether a program of research and further development of the technology could ultimately make it a technically feasible alternative to current open-pit and underground coal mining operations. ? What the nature and cost of further research and development work in the technology should be.

Jan 1, 1971

By H. P. Cole

Methods for increasing compliance to medical and health regimens were reviewed along with similar research in industrial and mine safety. Principles abstracted from both areas suggest that compliance with health and safety practices is improved when: (1) Workers are cooperatively involved with management in setting safety goals and planning for their achievement, (2) the program has the support of workers, supervisors, and management, (3) specific tasks responsible for injuries are targeted, (4) company organizational policy, educational, and behavioral programs work together toward the same goals, (5) workers and supervisors are frequently informed about their degree of correct task performance, and (6) programs are monitored by the frequency of correct performance of targeted tasks, (the means to lowered accident rates) rather than by lowered accident and injury rates themselves (the goals for safety programs.) Safety programs that incorporate these principles increase persons&apos; adherence to specified health and safety practices and subsequently decrease injury rates. It is recommended that: (1) a group of experts be convened to study the application of these principles to coal mine safety programs, and (2) a short course for safety trainers be developed to demonstrate how these principles can be used to strengthen existing mine safety programs.

Jan 1, 1985

By James M. Link

The first rule of mineral deposits always seems to be: the deposit is never close to the market. With gems and precious metals, this is not much of a problem. But with most of the mineral commodities used in today&apos;s world, transportation becomes a major part of the cost to the consumer. In Japan, for example, more than 60% of the cost of coal to the consumer may be attributed to transportation. The purpose of this article is to examine pipeline transportation, particularly mineral slurry pipelines, as a means of getting the mineral to market. Pipelines Today Regulated pipelines in the US totaled nearly 724 Mm (450,000 miles) in 1980 according to the Federal Energy Regulatory Commission (FERC). Worldwide, more than 210 Mm (125,000 miles) of new pipelines will start construction in 1982 at an estimated cost of more than $150 billion. According to the Bechtel Corp., the period between 1982 and 2000 will see major growth in pipelines of all kinds. In North America, major movements of crude oil from Alaska and synfuels from the Rocky Mountains will require the construction of nearly 16 Mm (10,000 miles) of new pipelines. Nearly 42 Mm (26,000 miles) of new gas transmission pipelines will be needed to tap both US and Canadian arctic gas fields and move the gas to major markets on both coasts. Other pipelines will be required to move additional quantities of natural gas from the Rocky Mountains to more populous regions. Bechtel says more than 24 Mm (15,000 miles) of slurry pipelines will be needed to transport coal from both eastern and western coal provinces to the Mississippi valley and coastal areas. In 1980, 125 companies delivered more than 1 km3 (6.5 billion bbl) of crude oil and 652 hm3 (4.1 billion bl) of products in the US. At the same time natural gas pipelines transported 498 km3 (17.6 trillion cu ft). The reported investment by these companies was nearly $20 billion at the end of 1980. Advantages of Pipeline Transportation The obvious success and vitality of the oil and gas pipeline industry is based, at least in part, on the fact that pipelines are a very efficient and low-cost method of transportation. This fact, coupled with the need for lower cost trans¬portation, has led to the marriage of oil and gas transmission technology and the bulk mineral solids transportation industry. The off-spring of this marriage is the mineral slurry pipeline. Slurry pipelines have a number of advantages over other transportation methods. One of these is the fact that they are buried-out-of-sight, out-of-mind. Second, they are relatively small users of labor because they lend themselves to automation and remote, or even computer control. Third, they offer an attractive economic alternative to other transportation systems. For example, for a 1000 km (621 miles) distance, rail costs of 1?/ km (1.7?/ton-mile) are about equivalent to slurry pipeline costs. But, as the distance increases, pipeline cost per t-km continues to drop while equivalent rail charges remain essentially insensitive to distance. Where existing rail, barge, or ocean ships are available, the cost of new construction associated with a slurry pipeline probably will render it noncompetitive. However, where new rail or other construction is needed, the cost of a slurry pipeline is very competitive. About three-fourths of the cost of a pipeline is in pipe, fittings, and construction. Nearly a fifth of the investment is in pump stations and the remainder is in right-of-way, surface facilities, utility acquisition, communication facilities, and other support areas. The fact that pipelines are capital intensive is, at the present time, a mixed blessing. For example, the delivered cost of coal in a hypothetical project doubles when the cost of money rises from 9% to 17.5% per year. The largest cost element for delivered coal in this example is in depreciation and finance charges. The next largest cost element is electricity to power the pumps, with labor making up the smallest increment of the delivered cost of coal. World Slurry Pipelines The idea of a slurry pipeline was probably first investigated in the latter 19th century, but the first ones were successfully built and operated in the US in the early 1900s. Today, they are fairly common worldwide. Table 1 shows a number of the world&apos;s slurry pipelines. An examination of these will further emphasize the fact that pipelines provide a cost competitive alternative where new construction of a transportation system is required. In each of the examples, the desired mineral commodity is located in a remote corner of the world, markets are at a great distance, value of the commodity is not great, and no other transport system is available.

Jan 10, 1982

By H. Lyn Bourne

This is the sixth annual tabulation that gives mineral royalty information. This tabulation, like the previous editions, includes all of the earlier data plus more than 100 new entries. In addition to the royalty information, some entries show the cost of purchased reserves. An article in the December 1984 Kentucky Coal Journal could have far reaching applications. It explains that a banking firm examined the "reserve risk" as one of six factors for analysis when financing a coal project. The firm concluded that coal reserves in Kentucky are worth an average of $0.29/t ($0.26 per st) in-place. The banking firm based this in-place value on its examination of more than 50 leases. In-place value is especially important when trying to establish a mineral value for a given deposit. Ernie Lehman wrote to this author after last year&apos;s column appeared and he quoted the lease rates for lead and zinc on federal land. He states, from first hand information, that the leases are based on a percentage of net smelter return as follows: 4% for the first five years, 4.5% for the next five years, and 5% thereafter. The figures used in last year&apos;s column ($ per st) came from dividing reported royalty income by reported production. The information reported last year merely translates the results of the net smelter return (as written in the lease) to royalty income per ton to the US Government for the various years. Statistics from the Minerals Management Service of the Federal Government show that oil and gas account for 40% and 53%, respectively, of the Federal royalty income. Coal accounts for about 4% and all other minerals, the remaining 3%. Oil and gas leases are mostly at 12.5%. Most of the coal leases are at 12.5% for surface mining operations and 8% for underground mining. Many of the other commodities are leased at 5%, which has been true for decades. Relative to Federal leases, this year&apos;s table shows some percentages and some costs per ton. This was done to give information about the leases and also to show income produced by various commodities. Note that some commodities have a variation in $/ton even though the royalty percent is the same. These variations result from a combination of factors that influence the value of a mineral commodity. The same caveats that have appeared in previous columns relative to factors that influence the value of reserves are restated. The data in the table serve as a general reference and are not intended to present an absolute scale of reserves costs. Several factors influence the value of any mineral commodity. The factors include: • location/transportation (especially significant for most industrial minerals); • market conditions both for the commodity and the company acquiring the right to mine it; *quality or grade of the deposit as it reflects the amount of processing/beneficiation necessary to produce the finished product; and • legislative restrictions and conditions for mining a given material at a given site. These and other factors will significantly influence the cost of reserves. The table shows the range of royalties and the variations in the basis of computing the fees. In a few instances, it gives the cost of acquired reserves based on purchase price and quantitative estimates of the material in place. The table lists the commodity in column 1, and the subsequent columns give the location - either by state, US geographic area or Canadian province; the cost - R for royalty and P for purchased cost; and the year in which the lease or purchase agreement occurred. The last column offers either comments or gives a footnote number for more information. The cost column requires more explanation. The units may refer to cubic yards (/yd), cubic feet (/cu ft), tons (/T, 2000 lbs) or acres (/A). 1 cu yd = 0.7646 m3 1 cu ft = 0.2832 m3 1st=0.9078t 1 acre = 0.4047 hm or m2 In several cases, the royalty is a percentage of the sales price of the finished product and is levied against either the amount produced or the amount sold. We thank those who contributed information for the table: G. Anderson, B. Brown, R. Ganis, K. Hauser, E. Lehman, and S. Sims. Their help is appreciated as are the contributions by those who asked to remain anonymous. This column requires additional information each year. That data must come from people in the minerals industry who have royalty information. What appears in the table is only a very small fraction of the cost information relative to the various commodities and the geographic areas for which there is royalty information. If you find this column of interest and possess information write the author at the address given in the author&apos;s box or call 313-462-0005.

Jan 7, 1988

By John S. Crandall

THE mineral garnet, while ordinarily considered a semiprecious gem stone or a second-grade industrial gem, has also proved itself in the field of industrial abrasives. Its use is well known as a sandpaper grain, and as a sandblasting sand its qualities are rapidly becoming recognized in more and more industries. Production of garnet as an abrasive is confined chiefly to two areas in the United States, North Creek, N. Y., where the Barton Mines Corp. operates, and Emerald Creek, Benewah County, Idaho, where Occurrence: Garnets in the Emerald Creek area occur as disseminated crystals in beds of micaceous schists of the Belt Series, which in this section are estimated to be close to 4000 ft thick. The schists are high in alumina and silica with iron, manganese, and magnesium. Subjection of the original sediments to high temperatures and pressures caused metamorphism to take place with the resultant re-crystallization of high alumina-silica minerals such as garnet, mainly spessartite and almandite varieties, cyanite, sillimanite, chlorite, actinolite, tourmaline, biotite, and muscovite, with minor amounts of ilmenite and magnetite. Quartz is also present in considerable amounts. Fast erosion of the soft mica schists on exposure to weathering has created extensive alluvial deposits containing up to 10 pct garnet having a maximum grain size of 3/16 in. These alluvial sands and gravels are now being treated for the recovery of garnet sands. Treatment: Overburden of 1 to 4 ft must be stripped to expose the garnetiferous gravels. This operation and the subsequent feeding of the gravels to a trommel-screen washing plant are performed by a % yd dragline. The trommel-screen openings are 3/16 in., thus allowing a separation and concentration based on grain size, since over 95 pct of total free garnets are minus 3/16 in. All plus 3/16-in. material is wasted at this point. The minus 3/16-in. material is further concentrated in a sand-drag classifier, where the slimes and silts are washed out and wasted. The sand product from the classifier varies in garnet content from 20 to 60 pct according to the particular section of ground being worked. This sand product is trucked to a jig plant where two sized fractions are made in a trommel-screen. The minus 3/16-in. plus 10-mesh portion is fed to a Pan-American two cell 42-in. jig. The minus 10-mesh portion is treated in a Bendelari three cell 42-in. jig. The jig concentrates are combined to form a 98 pct garnet sand. The jig tailings contain 3 to 5 pct garnet which is mainly flat crystals and chips which will not settle into the jig hutch. Subsequent treatment of these tailings in a scavenger jig followed by drying and electromagnetic separation will, according to tests, reduce the garnet losses in the tailings to something around 1 pct. Jig treatment of this feed approaches ideal as the major portion of the garnet crystals are the natural dodecahedrons and so are, in general, close to spherical. The specific gravity of pure garnet is 4.2, while the next heaviest mineral in the feed is cyanite with a specific gravity of 3.6, then quartz with specific gravity of 2.6. The garnet concentrate is practically free of quartz. The predominant impurity is cyanite which amounts to about 1.5 pct. The rod-like crystals of cyanite appear to up-end in the jig and go into the hutch with the garnets. Some ilmenite and magnetite appear in the concentrate but in very minor amounts. Subsequent washing in a sand-drag classifier removes fine silts and iron oxides. The gravel feed to the washing plant will average 8 pct recoverable garnet content. Concentration ratio in this plant runs about 2.5 to 1. Washing-plant concentrate as fed to the jigs will average 45 pct garnet by weight. Concentration ratio of jigging runs about 2.2 to 1. The garnet concentrate is dried in a rotary oil-fired drier and then fed to vibrating screens in closed circuit with crushing rolls. Practically any grit from 10-mesh down to 150-mesh grain size may be graded to specifications in two 3-deck vibrating screens. The present production, however, is approximately 75 pct No. 36, 15 pct No. 60. and the balance No. 80 and No. 100. Metal-screen cloth is used for sizes down to 36 mesh. From 36 mesh and finer, silk-screen cloth is used since it has less tendency to blind. All garnet sand is bagged in 100 lb self-sealing, sleeve-type paper bags. Practically all shipments are made in carload lots. Car loading is convenient since the plant is in Fernwood on the tracks of a branch line of the Milwaukee railroad. Truck shipments can and are made occasionally.

Jan 1, 1951

By John S. Crandall

THE mineral garnet, while ordinarily considered a semiprecious gem stone or a second-grade industrial gem, has also proved itself in the field of industrial abrasives. Its use is well known as a sandpaper grain, and as a sandblasting sand its qualities are rapidly becoming recognized in more and more industries. Production of garnet as an abrasive is confined chiefly to two areas in the United States, North Creek, N. Y., where the Barton Mines Corp. operates, and Emerald Creek, Benewah County, Idaho, where Occurrence: Garnets in the Emerald Creek area occur as disseminated crystals in beds of micaceous schists of the Belt Series, which in this section are estimated to be close to 4000 ft thick. The schists are high in alumina and silica with iron, manganese, and magnesium. Subjection of the original sediments to high temperatures and pressures caused metamorphism to take place with the resultant re-crystallization of high alumina-silica minerals such as garnet, mainly spessartite and almandite varieties, cyanite, sillimanite, chlorite, actinolite, tourmaline, biotite, and muscovite, with minor amounts of ilmenite and magnetite. Quartz is also present in considerable amounts. Fast erosion of the soft mica schists on exposure to weathering has created extensive alluvial deposits containing up to 10 pct garnet having a maximum grain size of 3/16 in. These alluvial sands and gravels are now being treated for the recovery of garnet sands. Treatment: Overburden of 1 to 4 ft must be stripped to expose the garnetiferous gravels. This operation and the subsequent feeding of the gravels to a trommel-screen washing plant are performed by a % yd dragline. The trommel-screen openings are 3/16 in., thus allowing a separation and concentration based on grain size, since over 95 pct of total free garnets are minus 3/16 in. All plus 3/16-in. material is wasted at this point. The minus 3/16-in. material is further concentrated in a sand-drag classifier, where the slimes and silts are washed out and wasted. The sand product from the classifier varies in garnet content from 20 to 60 pct according to the particular section of ground being worked. This sand product is trucked to a jig plant where two sized fractions are made in a trommel-screen. The minus 3/16-in. plus 10-mesh portion is fed to a Pan-American two cell 42-in. jig. The minus 10-mesh portion is treated in a Bendelari three cell 42-in. jig. The jig concentrates are combined to form a 98 pct garnet sand. The jig tailings contain 3 to 5 pct garnet which is mainly flat crystals and chips which will not settle into the jig hutch. Subsequent treatment of these tailings in a scavenger jig followed by drying and electromagnetic separation will, according to tests, reduce the garnet losses in the tailings to something around 1 pct. Jig treatment of this feed approaches ideal as the major portion of the garnet crystals are the natural dodecahedrons and so are, in general, close to spherical. The specific gravity of pure garnet is 4.2, while the next heaviest mineral in the feed is cyanite with a specific gravity of 3.6, then quartz with specific gravity of 2.6. The garnet concentrate is practically free of quartz. The predominant impurity is cyanite which amounts to about 1.5 pct. The rod-like crystals of cyanite appear to up-end in the jig and go into the hutch with the garnets. Some ilmenite and magnetite appear in the concentrate but in very minor amounts. Subsequent washing in a sand-drag classifier removes fine silts and iron oxides. The gravel feed to the washing plant will average 8 pct recoverable garnet content. Concentration ratio in this plant runs about 2.5 to 1. Washing-plant concentrate as fed to the jigs will average 45 pct garnet by weight. Concentration ratio of jigging runs about 2.2 to 1. The garnet concentrate is dried in a rotary oil-fired drier and then fed to vibrating screens in closed circuit with crushing rolls. Practically any grit from 10-mesh down to 150-mesh grain size may be graded to specifications in two 3-deck vibrating screens. The present production, however, is approximately 75 pct No. 36, 15 pct No. 60. and the balance No. 80 and No. 100. Metal-screen cloth is used for sizes down to 36 mesh. From 36 mesh and finer, silk-screen cloth is used since it has less tendency to blind. All garnet sand is bagged in 100 lb self-sealing, sleeve-type paper bags. Practically all shipments are made in carload lots. Car loading is convenient since the plant is in Fernwood on the tracks of a branch line of the Milwaukee railroad. Truck shipments can and are made occasionally.

Jan 1, 1951

By J. A. Rowland, C. W. Funk

THIS investigation is a part of the United States Bureau of Mines work in conserving the Nation&apos;s resources. The isothermal sections presented were developed as a guide to a comprehensive investigation of the properties and fabricating characteristics of copper-base alloys containing manganese and tin. These alloys are being investigated as possible replacements for the commercial bronzes containing substantial quantities of tin. Development of the isothermal sections has, therefore, been limited to the area between 0 to 20 pct Mn and 0 to 25 pct Sn. Since Heusler1 reported the ferromagnetic properties of the Cu-Mn-Sn alloys in 1903, the system has been the subject of several investigations. These studies were largely concerned with correlation of magnetic properties and crystal structure.2-6 Vero,&apos; however, established the solidus surface of the system, between 20 pct Mn and 40 pct Sn, by thermal and microscopic methods. The a, ß, and y solid solutions were also observed by Vero and reported to extend into the interior of the ternary system. For the present study, the annotated diagram of the Cu-Sn system prepared by Raynor&apos; and the Cu-Mn diagram by Dean and coworkers9 were used as the binary borders of the ternary system. Procedure The metals used in this investigation were electrolytic manganese, washed cathode copper, and three-star tin. The manganese, produced at the Boulder City pilot plant of the Bureau of Mines, had a purity of 99.9 pct; the copper contained less than 0.004 pct Bi and 0.001 pct S; the tin contained 0.070 pct Pb, 0.070 pct Sb, 0.050 pct Fe, and 0.065 pct Si. Heats of 1 lb were melted in alundum crucibles with a 3-kw induction furnace. The heats were chill-cast in iron or copper molds to form 6-in. ingots 3/4 in. in diameter. Molds were washed with graphite or zirconia, depending on the manganese content, and preheated to 150°C. The maximum impurity content of any alloy was 0.03 pct Fe, 0.02 pct Si, and 0.02 pct Al; in most cases, they contained less than half this quantity of any of these impurities. Composition of the alloys and their treatment prior to homogenizing are shown in Fig. 1. Designations adjacent to each composition refer to the heat numbers. With a few exceptions, ingots containing 20 pct Sn or less were hot-swaged. These ingots were swaged, after a 24-hr preheat at 650°C, to obtain a total reduction of 70 pct in cross section. Intermittent reheating was necessary, and reductions were limited to either 0.025 or 0.050 in. in diameter per pass. Selection of a 200-hr homogenizing treatment was based on extensive experiments which indicated that virtual equilibrium was reached after 150 hr. Structures were retained by quenching in water. Grain sizes in the specimens homogenized at 350" and 450°C were generally small. These homogenizing treatments were preceded by a 50-hr anneal at 650°C to facilitate phase identification by increasing the grain size in these specimens. The specimens were furnace-cooled from this temperature to the homogenizing temperature, and the heat treatment was continued for the standard 200-hr period. The homogenized samples were bisected to provide an internal section for metallographic examination. Optimum definition of the microstructure was obtained by successive polishing and etching. Quarter-strength ASTM copper reagent No. 13 gave the most satisfactory results.10 Filings from interior sections of the samples were used for diffraction studies. These filings were annealed in evacuated glass tubes at homogenizing temperatures for periods exceeding 50 hr and quenched in water. During this treatment the manganese content was reduced by significant amounts, probably by vaporization. This change in composition was considered in applying the diffraction data, which were used only as a means of identifying phases where the data were consistent with the metallographic evidence. Phase Identification The a solid solution is readily distinguished by the copper-colored, twinned, polyhedral grains, as shown in Fig. 2, and a face-centered cubic X-ray pattern. Tukon indentations, using a 25-g load, indicate a Knoop hardness of 155 for a quenched from 650°C. Using white light developed by a Wratten 78 A filter, the ß grains in etched specimens have a dark brown tint in contrast to the lighter copper-colored grains of the a phase. Knoop hardness values of 284 were obtained for the ß structure in specimens quenched from 650°C. The ß structure was also distinguished from the a phase by its body-centered cubic lattice. The acicular structure appearing in Fig. 3 was evident in several alloys of higher tin content after quenching from either 750&apos; or 650°C. Since a trans-

Jan 1, 1954

By Thomas E. Wayland

AN isopachous map is one on which lines connect points of equal thickness of a given unit. This type of map is used by the Florida Phosphate Project of the U. S. Geological Survey to represent the economic phosphate deposits known as matrix and the waste material, or overburden, that overlies the matrix. The top of the bed on which the phosphate was deposited is known as the basement and a subsurface contour map of this old buried erosion surface is known as a basement map. Recent experiments have been made in preparing maps that show tonnages and grades of the phosphate content of the matrix. Few of the operating companies in the Florida phosphate district have applied isopachous (Greek isos, equal and pachys, thick) to mapping. The writer believes there is a need for the techniques discussed herein and that they can be applied to mapping other geologically similar areas in either economic or scientific investigations. The land-pebble phosphate district of Florida occupies a compact area in the west-central part of the state. It includes mainly the following land survey divisions: Ts. 27 S. through 32 S. and Rs. 20 E. through 26 E. The town of Mulberry, Fla., is in the approximate center of the district. The strata of the area, which is part of the Gulf Coastal Plain, occur in thin formations with broad outcrop belts, and low dips. The topography is subdued and gently rolling with three marine terraces, which are found at 30, 100, and 150 ft above sea level,&apos; accounting for most of the relief. Occasional small sinkhole lakes are present, most of them above the 150-ft shoreline. The phosphate deposits occur in unconsolidated sediments such as clays, sands, and sandy clays. They are overlain by a heterogeneous assemblage of sands, clays, muck, and iron-cemented sand, easily penetrated, in most cases, by a hand auger or drill. Limestone, locally called bedrock, or a calcareous bedclay, thought to be a residue of the limestone, directly underlies the phosphate deposits. General Requirements Most companies and independent prospectors operating in the district have furnished prospecting data to the U. S. Geological Survey. The information is recorded on either field logs or prospecting maps and includes the following information for each hole drilled: location of the hole, thickness of the overburden, thickness of the matrix, phosphate content in long tons per acre, grade of the phosphate content expressed as the percentage of bone phosphate of lime (P2O5 x 2.18) or BPL, and the per- centages of iron-aluminum oxides and insolubles. The phosphate is classified according to size as either pebble or flotation material. The milling processes of the companies vary, and the size classification is necessarily different in many cases. However, pebble may be considered as larger than 14 mesh and flotation material as smaller than 14, but larger than 150 mesh. Some prospecting data include the exact depth at which bedrock or bedclay was reached, and these figures greatly increase the reliability of the data both for isopachous mapping and for mapping the basement. A drilling density of four holes per 40 acres of land furnishes a minimum amount of data for isopachous and related mapping. From the minimum of four, densities up to 32 holes per 40 acres are used. The various drilling densities may influence the choice of the proper scale. Selection of the proper scale is dependent upon the known drilling densities, the subsurface variations to be shown, the extent of the area to be mapped, and the detail desired in the completed map. Scales of 1:24,000, 1:4800, and 1:2400 are used in isopachous and related mapping by the Florida Phosphate Project. The 1:24,000 scale is used most effectively with drilling densities not exceeding eight holes per 40 acres. The subsurface variations should be relatively low and uniform, permitting the use of smaller intervals without undue crowding of the lines. Comparatively large areas can be mapped on this scale, but minute detail is necessarily sacrificed, because the information is drawn from a maximum drilling density of only eight holes per 40 acres. Isopachous and related maps of the 1:4800 scale are made of areas on which the drilling information covers from 4 to 16 holes per 40 acres. Moderate subsurface variations with relatively sharp gradations can be shown accurately. The area represented by the maps is reduced considerably in favor of detail. The 1:2400 scale is most frequently used by the Florida Phosphate Project. It lends itself particularly well to isopachous and related mapping, being easily adapted to the multiform drilling data available. Maps of this scale are prepared with information ranging from 4 to 32 holes per 40 acres; however, use of the minimum drilling density on the

Jan 1, 1952

By E. R. Pantos

US coal production increased 3% to a record 832 Mt (917 million st) in 1987, according to preliminary Energy Information Administration (EIA) data. Most of the increase was due to the unusually warm summer and the increased coal-fired electricity demand for air conditioning. Drought conditions in the Northwest lowered hydroelectric generation, contributing to the higher use of coal by electric utilities. However, excess production capacity and increased competition from natural gas added further downward pressure on delivered coal prices, continuing the declining price trend of the last several years. Coal production by region Appalachian coal production reached 402 Mt (443 million st) in 1987. This is about 3% more than in 1986 and just short of its 1984 peak. Virginia produced 7% more than its 1986 level, due to increased coal-based electricity demand. Smaller percentage increases were attained in eastern Kentucky, Pennsylvania, and West Virginia. Ohio and Alabama were the only major coal-producing states in Appalachia that produced less coal in 1987 than in the previous year. In the Interior region, 1987 coal production reached a record 182 Mt (201 million st), about 2% above 1986. Illinois is the largest coal-producing state in that region. Production there fell about 3.7%, primarily because of lowered demand at electric utility plants in Alabama and Tennessee. However, this was more than offset by production increases in western Kentucky, Texas, and Indiana. After falling slightly in 1986, coal production in the West resumed its upward trend to a record 248 Mt (273 million st). Wyoming is that region&apos;s dominant coal producer. It mined 7% more coal than in 1986. Wyoming&apos;s increase offset the 11% decline in New Mexico production and smaller declines in other Western states. Coal consumption and prices Domestic coal consumption reached a record 759 Mt (837 million st) in 1987. Most of the increase resulted from the unusually warm summer that led to higher than normal peak demands for electricity in the third quarter. Electric utility coal consumption was also boosted by the below-normal hydroelectric generation that resulted from low water levels in the Northwest. These factors more than offset the effects of increased nuclear power generation and competition from natural gas. As a result, coal provided 56.9% of electric utility generation in 1987, compared with 55.7% in 1986. Electric utilities consumed a record 651 Mt (718 million st) of coal. While electric utility coal consumption increased, the average cost of coal to utilities was down, from $1.58 per million Btu in 1986 to $1.52 in 1987. The average price of coal delivered to utilities in 1987 was $35.09/t ($31.84 per st), the lowest since 1981. Coal prices were held down by a glut of coal supplies and by a decline in the average price of natural gas to electric utilities. Coal consumption at coke plants in 1987 was 2.6% above the 1986&apos;s 32.5 Mt (36 million st). This increase reflects steel&apos;s recent resurgence. Coking coal demand, though, was reduced by the increased use of continuous casting and electric furnace technologies. The average price of coal delivered to coke plants fell from $56.02 to $51.35/t ($50.83 to $46.59 per st). Coal consumption in the industries other than coke plants amounted to 68.2 Mt (75.2 million st) in 1987, slightly less than in 1986. Coal lost part of its market share to natural gas, which was priced favorably relative to coal. The average price of coal delivered to manufacturing plants was at its lowest level of the current decade. It fell from $39.50/t in 1986 to $37.27/t ($35.84 to $33.81 per st). Residential and commercial coal consumption fell to 6.2 Mt (6.9 million st) that is 9.8% below 1986 and the lowest since 1980. This decline is also attributed to the relatively low price of natural gas in 1987. Coal stocks Coal consumers held 169 Mt (186 million st) of coal stocks at the end of 1987, 5.9% above the 1986 year-end level. Most of this increase was due to stockpiling by electric utilities in anticipation of a possible strike by the United Mine Workers of America early in 1988. Stocks at electric utilities increased to an estimated 154 Mt (171 million st) at the end of 1987, 5.6% higher than the previous year. This was much lower than 1984&apos;s stock buildup. Coke plants increased their coal stocks by 30% in 1987 to 3.5 Mt (3.9 million st), rebounding from the unusually low inventories of 1985 and 1986. Other industries increased their stocks slightly in 1987. Coal producers and distributors ended 1987 with 25 Mt (28 million

Jan 5, 1988

By K. P. Gupta, A. K. Pal, L. Chandrasekaran, U. P. Singh

The Mn-Sn binary system, investigated at the high-manganese end and between 500° and 1000° C, shows four phases at temperatures below 727"C, namely the u Mn, the p Mn, the Mn3 Sn, and the Mn, Sn phases, while at higher temperatures only the last three phases remain stable. The solubility of tin in a Mn is very small and the maximum solubility of tin in P Mn phase appears to be abollt 10 at. pct Sn. The solubility range of the Mn,Sn and the Mn, Sn phases is 23.0 to 26.0 and 37.0 to 40.0 at. pct Sn, respectively. The lattice parameter of the 13 Mn phase increases with increasing tin content. The Mn, Sn thase is hexagonal with and appears to be basically of the Ni3 Sn type structure except that for quite a few X-ray diffraction lines the calculated and observed relative intensities do not agree well. The Mn-Sn binary system has been studied by several investigators.&apos;&apos;~ Their results indicate that three intermediate phases, namely, the Mn3Sn, Mn,Sn, and MnSn, phases, exist at high concentrations of tin. However, so far the proper phase equilibrium has not been established at the manganese-rich end and very little data is available for the composition range between pure manganese and Mn3Sn. Moreover, earlier investigators differ in their opinion about the exact composition range at which the Mn3Sn and Mn,Sn phases appear and some doubt has been cast by some investigators3 regarding the structure of Mn3Sn phase which has been reported to be isotypic with Ni3Sn (Mgscd type) structure. In this investigation an attempt has been made to establish the proper phase equilibrium between pure manganese and Mn + 50 at. pct Sn composition in the temperature range of 500" to 1000°C. PROCEDURE The raw materials used were from three different sources. For exploratory work five alloys containing 5, 10, 15, 20, and 25 at. pct Sn were prepared using 99.9 pct pure manganese and tin supplied by E. Merck & Co., Germany. The rest of the alloys and one more 25 at. pct Sn alloy were prepared using 99.9 pct Mn supplied by Gallard Schlesinger Chemical Mfg. Corp., U.S.A., and 99.999 pct Sn supplied by Semi Elements Inc., U.S.A. Weighed amounts of manganese and tin were melted in recrystallized alumina crucibles in an inert gas (argon) high-frequency induction melting furnace. By careful control of temperature and time of melting the losses were reduced to below 0.2 pct in all cases. Since the losses were very small, no attempt was made here to analyze the samples chemically. Alloys were wrapped in molybdenum foil and sealed in small evacuated fused silica capsules. The alloys were annealed at different temperatures, controlled within + l°C, for sufficiently long periods to attain proper phase equilibrium, and subsequently quenched in cold tap water. The annealing periods used at different temperatures were 15 days at 500°C, 7 days at 600°C, 5 days at 700" and 750°C, 3 days at 800°, 850°, and 900°C, 2 days at 968"C, and two alloys annealed at 968°C were reannealed for 10 hr at 1000° C. From each annealed specimen, a part was utilized for metallographic study while another piece was used for X-ray diffraction study. 1.0 pct HNO, solution and oxalic acid solutions of concentrations 0.05 to 1.0 pct were used for etching Mn-Sn alloys above and below the MnsSn composition, respectively. Since all alloys were brittle, X-ray specimens were prepared using the as-crushed -325 mesh alloy powders. Only one 25 at. pct Sn alloy powder was reannealed in an evacuated silica capsule at 800°C for 5 min and water-quenched. X-ray diffraction patterns . for the Mn3Sn phase with the as-crushed and the reannealed powders did not show appreciable change. Un-filtered iron radiation at 25 kv, 15 ma was used with either Norelco 114.6-mm-diam Debye Scherrer Camera (for phase identification) or Norelco 12-cm-diam symmetrical focusing camera (for lattice parameter determination of the 6 Mn phase). The estimated accuracy of lattice parameter determination for the focusing camera was * 0.001A. RESULTS AND DISCUSSIONS The results of metallographic and X-ray diffraction study made with different alloys are shown in Fig. 1. The variations in lattice parameter with composition for the p Mn and the Mn3Sn phases are given in Tables I and 11, respectively, and the lattice parameter as a function of composition for the p Mn phase is shown in Fig. 2. The lattice parameter of the p Mn phase increases with increasing tin content while for the Mn,Sn phase the data obtained from two two-phase alloys and one single-phase alloy indicate increase in a,, and decrease in c, parameters with increasing tin content. The results, Fig. 1, indicate that the solubility of

Jan 1, 1969