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By F. Chiesa, Jeremy Carignan, B. Duchesne, D. Levasseur

"Aluminium A206 (AlCu4) is one of the strongest aluminium casting alloys with tensile properties approaching those of ferritic ductile iron. If the specification on the silicon content (< 0.05%) is easy to achieve during alloy ingot production, it is not the case in a foundry, where returns are used which are not perfectly cleaned from the silicate containing molding sand (Olivine). In alloy A206, the Si content is controlled in order to avoid the formation of a low melting point intermetallic phase (such as phase Q Al5Cu2Si6Mg2); these phases will melt during the solutionizing heat treatment to which alloy A206 must be subjected to develop its outstanding mechanical properties. This incipient melting may deteriorate the mechanical properties of Alloy A206 if the solutionizing treatment is applied at the standard temperature of 530°C.In the present study, a lower temperature solutionizing (520°C) was applied to a high Si content 206 alloy in order to reduce or eliminate the possible incipient melting. This was carried out on a complex shape industrial casting weighing 14 kg (31 lb) and 560x410x250 mm (22x16x10”) in size poured in aluminum 206 with a high silicon content (0.22% vs 0.05% max for alloy A206). Twenty ASTM E8 subsize tensile specimens were excised from two locations in the casting in order to compare the tensile properties for the two solutionizing temperature used (530 and 520°C). Optical and SEM microscopy allowed to investigate the causes of the differences in the properties observed between the 2 zones of the casting studied for 3 different heat treatments. INTRODUCTIONIn spite of their “textbook” outstanding mechanical properties (Up to twice the strength and elongation of the widely used A356 alloy), aluminum-copper alloys of the 2xx series are seldom used because of their propensity for hot tearing [Li, 2011], the segregation of copper [Cho, 2003], and their sensitivity to stress corrosion cracking [Jean, 2009, p. 113)]. Also, because of their wide freezing range and high thermal conductivity, a wide mushy zone makes feeding difficult, unless strong temperature gradients are created by liberal use of chills and risers in the critical zones of the part. Because they are prone to hot tearing, grain refining of Al-Cu alloys is particularly important, titanium playing a major role in the process [Sigworth, 2003]; grain refining has also been shown to be beneficial to fatigue strength [Jean, 2009, p. 103]. The practical basis of foundry practices pertaining to these alloys is outlined in an AFS webinar [Weiss, 2014]."

Jan 1, 2016

By O. M. Bishop

From October 1943 to March 1944, the Bureau of Mines conducted core drilling at the Victory fluorspar mine, Hardin County, Ill. to test the Fredonia limestone and the included sub-Rosiclare sandstone zone beneath the existing stopes. Twenty-one vertical holes were drilled to an average depth of 60 feet. Ore was penetrated in six holes. Since its discovery in 1927, the Victory mine has produced approximately 5 percent of the total fluorspar mined in the United states after that date. The ore discovered in the early mine development is now exhausted, except that contained in pillars drifting to the northeast has partly developed one substantial ore body, and company drilling has indicated another nearby. The increasing need for fluorite for war use in 1943 suggested the desirability of mining the pillars, provided it could be established that no ore beds existed below the stopes or in areas to which access would be obtained through the present workings. This work describes the work accomplished during the project and presents the data obtained.

Jan 1, 1948

By C. S. Parsons

Custom milling-plants have been operated successfully for a great many years on this continent. One of the outstanding examples in the United States is the Golden Cycle mill at Colorado Springs, which mills the gold-telluride ores from the Cripple Creek district. In eastern Canada, the Northern Custom Concentrators, Limited, at Cobalt, successfully treated the cobalt-silver ores of that area for a number of years. Flotation plants have entered the custom field in the past few years. The building of flotation concentrators to treat ores on a custom basis indicates the advance which has been made in flotation. In fact, three years ago it was considered almost impossible to treat successfully a mixture of miscellaneous ores by flotation. The methods used at that time were considered so delicate that any change, even in a single ore, would usually result in decreasing the effectiveness of the separation. Today we find four large custom flotation plants in the United States and one in Canada for treating complex lead-zinc ores from many different localities. The advent of these mills was brought about by the serious problem confronting the lead-smelting plants due to increasing amounts of zinc in the ores presented for treatment. It was necessary to penalize the mines shipping high-zinc ores. The penalty is justified, because high zinc in the lead blast-furnace charge not only decreases the tonnage of the furnace but increases the amount of flux necessary, as well as increasing the metal loss in the slag. It has been estimated that in the Salt Lake valley alone, 75,000,000 pounds of zinc was lost in the slag-dumps from the operations of three large lead blast-furnace plants. It has also been estimated that $150,000 worth of additional flux was required to slag this zinc. It was to avoid at least a part of this wste and to help maintain an ore supply that custom flotation concentrators were built by the smelters.

Jan 1, 1928

By H. C. Rickaby

The Bannockburn gold area covers the four townships of Bannockburn, Argyle, Hincks, and Montrose, lying in the western part of the Matachewan district of Ontario (see Figure 1). The geology of these four town-ships was studied by H. C. Cooke of the Geological Survey, in 1917-18, and is shown in his map of the district. In the fall of 1930, a rich gold-bearing quartz vein was discovered in the northwest part of Bannockburn by B. S. Ashley. A prospecting rush followed the discovery; many claims were staked in the Bannockburn area, and, during the summer of 1931, a considerable amount of prospecting was done.

Jan 1, 1932

By S. E. Malouf, R. Hinse

"The Mineral possibilities of the Chibougamau area were first recognized by Richardson in 1870 and Low in 1885, both of the Geological Survey of Canada. The first record of interest by the Quebec Bureau of Mines is in 1904, when J. Obalski reported on the discovery of copper and asbestos by Peter McKenzie in the previous year. Reports of 1906 and 1907 covered the McKenzie Gold Mine find and other deposits. In 1908 Prof. Dulieux described the quartz vein on Merrill Island and the mineralized zone on the near-by mainland. In 1910 the Chibougamau Railway Commission, though commenting favourably on the possibilities of the area, concluded that a railway was not then warranted.In spite of this, prospecting in the district continued. The main Campbell Chibougamau deposit, the Blake showing on Merrill Island in Dore Lake, was found in 1920, and other discoveries were made soon after. Between 1927 the work of Mawdsley, Retty, and Norman tor the Geological Survey and the Quebec Bureau of Mines stimulated further interest. In 1934 Consolidated Chibougamau Goldfields Ltd., a subsidiary of Consolidated Mining and Smelting Co. of Canada Ltd., acquired the Merrill Island and Cedar Bay holdings of Chibougamau Prospectors Ltd. They embarked on an aggressive campaign, directed essentially at the gold quartz veins at Cedar Bay. Exploration there included a 522-foot shaft, 4,732 feet of drifts and crosscuts, and about 34,000 feet of surface and underground diamond drilling. Thirteen holes were also drilled on the replacement sulphide lens at Merrill Island. Operations were suspended in 1936, as the occurrence was too low grade for development at the existing metal prices and under difficult operating conditions.The completion of the Quebec Bureau of Mines highway from St. Felicien to Chibougamau in 1950 was followed by a project of detailed geological mapping by the Bureau. It also encouraged Campbell Chibou-gamau Mines to acquire the assets of Consolidated Chibougamau Gold-fields and commence a program of surface exploration. Between August 1950 and April 1951, 41,183 feet of diamond drilling was done on the Cedar Bay and Merrill Island holdings. The Blake showing off the west shore of Merrill Island developed at depth into a major orebody. A lease was obtained on adjacent ground from Merrill Island Corporation, and underground work commenced in November 1951. Development at Merrill Island to June 1st, 1955, included a four-compartment shaft to 1,225 feet, with 8 levels at 150-foot intervals, 5,317 feet of raises, 16,709 feet of drifts and crosscuts, 52,514 feet of surface and 96,593 feet of underground diamond drilling."

Jan 1, 1949

By Paul W. Jeran

Subsidence was monitored by the U.S. Bureau of Mines over the ends of longwall panels operating in the Pittsburgh, Kittanning, and No. 2 Gas Coalbeds of the northern Appalachian Coal Basin. The final subsidence over the finishing ends of three panels in the Pittsburgh Coalbed are compared with the subsidence measured over the rib at these panels. The characteristics of subsidence are different. At Mine A, data over the start of a longwall panel shows similar characteristics to the subsidence measured over the rib. Subsidence over the fmishing ends of panels in the Kittanning and No.2 Gas Coalbeds are also different from the subsidence over the rib. The use of a subsidence prediction model based on data gathered over the rib of a panel will not yield accurate results if it is applied to the finishing end of a longwall panel. Acceptable results may be obtained along the centerline over the starting end of a panel.

Jan 1, 1991

By R. Hogg

A standard procedure for determining size-specific gravity distributions of fine coal (including the -400 mesh fractions) by sieving and centrifugal sink-float separation is described. The significance of using a centrifugal sink-float technique and the effect of the magnitude of the centrifugal force on the separation of fine coal are confirmed in the present study. The relative merits of cumulative versus incremental sink-float are discussed. The procedure is evaluated quantitatively using data obtained for the products of froth flotation of -100 mesh coal.

Jan 1, 1994

By Tong Guan, Guangqin Fu, Lihong Deng, Yuanlin Chen, Xiaotong Zhou

"In natural world, some kinds of scheelite minerals are rich in molybdenum. Molybdenum exists in the crystal lattice of scheelite in the form of isomorphism, and the content of MoO3 varies from 2% to 10%. As a result, the flotability of seyrigite (scheelite with molybdenum) turns to be weaker. One seyrigite mine, which has been built for 3 years,still could not produce qualified seyrigite concentrate due to the weak flotability of seyrigite (Mo-bearing scheelite) in raw ore. In our experimental research, through adopting new seyrigite collector TA-3 and advanced mineral processing technology, the seyrigite concentrate with a grade of 65.62% WO3 at a recovery of 87.47% was obtained for the raw ore containing 0.19% WO3 and 0.074% Mo. By applying the new technical achievement, for the raw ore containing 0.272% WO3, the seyrigite concentrate assaying 68.03% WO3 with a recovery of 77.56% was obtained in industrial production.INTRODUCTIONIn natural world, some kinds of scheelite minerals are rich in molybdenum. Molybdenum exists in the crystal lattice of scheelite in the form of isomorphism, and the scheelite is changed to be seyrigite (Ca[(W,Mo)O4]) (Ma, Bian, & Wang, 2013). The content of MoO3 in seyrigite varies from 2% to 10% usually, the content of MoO3 in some of the seyrigite even reaches 34%. The flotability of seyrigite is weak (Zhang, 2007), and the seyrigite is difficult to be recovered by flotation with common collectors and regulators. One scheelite mine in Anhui China, which is a skarn type seyrigite mine, has not produced qualified concentrate yet since 3 years it has been built due to the limit of beneficiation technology. Therefore, it is necessary to conduct research on the collectors with high efficiency and developing advanced processing flow for this kind of ore."

Jan 1, 2016

By H. L. Bourne

The sand dunes which line the east shore of Lake Michigan resulted after the last, or Wisconsin, glacial period in Michigan. Melt-water streams carried vast quantities of glacial debris to the near shore areas of predecessors to Lake Michigan. As the lake levels receded, these vast quantities of sand were exposed to the prevailing westerly winds, which was the last major process that acted upon the sand and resulted in the formation of the dunes. 1 These dunes are one of the significant natural resources within the state. Michigan dune sands supply about 90% of the foundry sand used in the automotive industry, and 40% of the foundry sand used 2 for all purposes within the United States. The deposits owe their importance to the following factors: grain size distribution, grain shape, mineral composition and location with respect to the foundry sand market. In spite of their economic significance these I deposits are in jeopardy through legislative preemption.

Jan 1, 1975

By R. W. Bartlett

The extensive central Idaho black sand placer deposits are a potential source of rare earth elements, particularly the heavy rare earth elements that are not prevalent in commercial bastnasite deposits. These placers derive from the Idaho batholith. The geology, occurrence, and extent of these deposits are reviewed and the diverse mineral assemblages and complex mineral chemical compositions of these sands are described. Niobium, yttrium and thorium are also present in significant concentrations. The results of prior mining practice, including concentration and separation of the heavy minerals are reviewed. Technical approaches to extraction of rare earths from Idaho black sand concentrates are discussed.

Jan 1, 1992

By Luis Alberto Sfalsin Passos

Com o passar dos anos observa-se uma mudança nas características mineralógicas dos minérios de ferro lavrados nas novas minas da Samarco Mineração SA. De acordo com levantamentos podem ser constatadas mudanças nos percentuais da hematita porosa, hematita especular e da goethita na constituição do concentrado de minério de ferro. Pode-se verificar uma tendência de aumento do percentual de goethita bem como do PPC (perda por calcinação) característica intrínseca ao minério. Estes fatos implicam em uma maior dificuldade da concentração em garantir a qualidade da polpa de minério bombeada para as usinas de pelotização, visto que fatores como a granulometria do minério (partículas menores que 325#) e a superfície específica são afetadas pela mineralogia. Por consequência, há uma maior dificuldade no processo de filtragem para obtenção do pellet feed com a faixa de umidade necessária para o processo de pelotamento e queima. De outra forma, o aumento do PPC dificulta o processo de secagem das pelotas no forno de pelotização, exigindo um maior aporte térmico para retirada da água residual contida nos cristais que não evaporam a 100°C. Todos estes fatores contribuem para a qualidade da pelota queimada, interferindo na sua resistência física como compressão a frio e tamboramento. Esta pesquisa tem como objetivo apontar as variáveis que influenciam significativamente o tamboramento, para propor ações que melhorem o controle do processo produtivo de forma a buscar o ponto de equilíbrio entre qualidade e produção.

Aug 16, 2017

By Hongxu Li, Zhaobo Liu, Zihan Zhao

A completely new process of selectively recovering Sc3+ without impurity of Fe3+ from Bayer red mud was developed. Prior to water leaching process, red mud needs to be digested and sulfated at a low temperature and subsequently roasted at a higher temperature for a period of time. During the roasting process, the products such as SO2, SO3 and H2SO4 molecules of thermal decomposition of sulfites including NaFe(SO4)2 and NaAl(SO4)2 could be recycled and reused as the raw materials of sulphuric acid. The leaching results indicate that roasting temperature and roasting time are the main constraint on selective leaching performance of scandium. As the sulfated red mud sample is roasted at 750 degrees C for 40 min and subjected to water leaching at 50 degrees C for 30 min, 53.0 wt% Sc3+, <1.0 wt% Fe3+, <0.3 wt% Si4+ and ~8.9 wt% Al3+ could be leaching out. Of note, the solid-liquid separation process is quite easy to be carried out.

Mar 1, 2017

By M. R. Goldick

The Roan Antelope concentrator was originally designed with a nominal milling capacity of 6000 tons of copper ore per day but this was subsequently considerably exceeded. In broad outline the plant consisted of two No. 30 McCully gyratory crushers for coarse crushing, five 5½-ft Standard Symons crushers in open circuit for secondary crushing, ten No. 98 Marcy ball mills arranged in five sections, each two-mill section with primary and secondary Dorr classifiers for two-stage grinding. One stage of flotation roughing was followed by two stages of cleaning, all in airlift type pneumatic flotation machines, with the final concentrates filtered on Oliver drum filters. Tailing was thickened in a 250-ft Dorr traction type thickener before pumping to the tailing dam. Concentrating operations commenced in June 1931, on the completion of the first two ball-milling sections and output was increased during the next few months as the remaining sections were finished. From the start, very few technical difficulties were experienced but a number of evolutionary changes have been made as a result of operating experience and the availability of improved equipment. These have resulted in plant and ore-flow alterations although the general scheme of ore treatment remains substantially as before. The main changes to date were the installation of a new crushing plant at Storke shaft, some 7000 ft west of the original Beatty shaft; the completion of a two-stage secondary crushing plant; an increase of 20 pct in ball-milling capacity; a change-over from pneumatic to a mcchanical type of flotation machine for flotation roughing; the reduction of flotation cleaning from two stages to one stage. Ore Treated Typically the ore treated consists of metamorphosed shales and sandstones with shales as the principal mineral-bearing rock carrying chalcocite, bornite and chalcopy-rite as the main copper minerals. Chalcocite predominated in the original eastern end of the ore body but there is a pronounced trend to an increased proportion of the copper-iron sulphides as mining proceeds westward. Although the ore can be described as soft, this advantage is offset by the extremely fine dissemination of the copper minerals which calls for a high degree of comminution of the ore (go pct minus-zoo mesh Tyler standard) to give a reasonably satisfactory separation of mineral from gangue. Latterly, there has been a tendency for the ore to become harder as shown by a reduction in the ball-milling rate from a previous figure of over goo tons per ball-mill day to the present output of some 800 tons. Chalcopyrite is not so finely disseminated and has also proved more amenable to flotation than chalcocite. Nonsulphide minerals are mainly the so-called oxides such as malachite, cuprite and melaconite with some chrysocolla and

Jan 1, 1949

In 2005, CANMET-MMSL used its pool of expertise in ground control, open pit and underground mine engineering, mineral processing, information technology and applied mineralogy to launch a mine-to-mill research project. The objective of the project was to optimize the distribution of rock breakage energy among blasting, crushing and grinding activities, thereby maximizing the efficiency of energy utilization within comminution processes. Tests were designed to demonstrate that this approach could reduce energy consumption, thus contributing to the reduction of green house gas (GHG) production in the Canadian mining industry. The project was accomplished in collaboration with DynoConsult and Dyno Nobel Canada Inc., a world leader in commercial explosives. Data analysis from 12 blasts (six with electronic and six with pyrotechnic detonators) in three different geological formations of a quarry (soft, medium and hard rock) shows that electronic blasting technology increases rock fragmentation by 15% (in hard rock) to 20% (in soft rock). A more uniform rock fragment size distribution is also obtained. The impact of the improved quality blast product on downstream processes of the quarry is a productivity increase of 5% in rock excavation and transportation, an energy reduction of 10% at the primary crusher and an increase of 15% of the primary crusher throughput. Extended to the Canadian mining industry, such results would represent a GHG emission reduction of about 0.05 Mt/y. A soft-sensor prototype, based on primary crusher data (i.e., energy consumption, throughput, feed and product size distribution) and a simplified computerized process model, was developed for on-line measurement of the rock operating work index. Additionally, 385 rock samples and specimens were collected and prepared to determine the tensile and compressive strength of the rock material, to evaluate the aggregate crushing work index and to examine the development of micro- and macro-fractures by means of binocular and SEM microscopy. Blast-induced rock damage was evaluated at the end of the laboratory-testing program. Even though a reduction of rock hardness of around 30% was observed after blasting, the impact of the blasting method (i.e., use of electronic or pyrotechnic detonators) on the induced rock damage could not be observed. It is either absent or not perceptible with the accuracy of the measuring techniques and the number of samples analyzed.

Jan 1, 2008

By George McArthur

GEOLOGY One of the largest copper resources in North America occurs in the Bingham mining district, located 40 km (25 miles) southwest of Salt Lake City, UT. Anaconda&apos;s present contribution to the development of this resource is the Carr Fork mine. (Fig. 1, Bingham district plan map.) The Carr Fork story began about 300 million years ago when a thick sequence of sedimentary quartzite and limestone was deposited in the Oquirrh Basin. Subsequent earth stresses formed a series of west-northwest-trending overturned anticlines and synclines that were later offset by northwest-striking thrust faults and northeast-trending normal faults. (Fig. 2, Bingham district cross section.) About 38 million years ago the Bingham stock was intruded into some of the folds, resulting in the formation of porphyry ore in the intrusive, and fissure, replacement, and skarn ore deposits in the sediments. The Carr Fork skarns were formed preferentially in two limestone units, called the Yampa and Highland Boy, which average about 46 m (150 ft) in thickness, separated by about 76 m (250 ft) of quartzite. Copper mineralization consists almost entirely of chalcopyrite associated with pyrite and molybdenite in a gangue of garnet, diopside, magnetite, quartz, and carbonate. Molybdenum, gold, and silver are economic byproducts. (Figs. 3 and 4, indicated ore in plan and section.) EXPLORATION HISTORY The West Mountain mining district, of which Carr Fork is a part, was organized in 1863, after leaf-silver ore was discovered in the Carr Fork of Bingham Canyon. However, it was not until 1873, when the Bingham and Camp Floyd railroad to the town of Bingham was completed, that the mining camp flourished. Many small mining companies in the district were reduced, often through consolidation. By 1937 only the National Tunnel and Mines Co. was active in the Carr Fork area. In 1948 it declared bankruptcy, and was later purchased by The Anaconda Co. Anaconda conducted geologic investigations in the Bingham district until 1968, when a concerted diamond drilling program was initiated to fully evaluate the geologic potential of the Carr Fork area. The exploration program commenced with one diamond drill, but was enlarged to three drills following discovery of the first major ore intercept. Near completion of the exploration effort, large oil well rotary drills were used to drill to depths of 762 to 914 m (2500 to 3000 ft), after which the holes were completed by standard diamond drilling techniques. All holes were collared vertically, but considerable deviation occurred because of structure and rock conditions. Wedging and directional drilling were used to keep the holes on target and to provide multiple ore intercepts in the deeper holes. Geophysical surveys, in-

Jan 1, 1982

By Fred C. Bond, Walter Maxson

Since the last publication of tabulated results of grindability tests by the authors1 the total number of ball-mill tests made has more than doubled, and rod-mill tests have become increasingly important. Nearly all of the standard closed-circuit ball-mill and rod-mill tests made to date are included in the present tabulation, together with additional information not previously published. Standard Ball-mill Grindability Tests The method of conducting these closed-circuit tests has not been altered. A sample of the ore or other material is stage-crushed in rolls set at 1/16-in. opening with a 6-mesh screen until all of it has passed the screen. The combined minus 6-mesh screen undersize is mixed, sampled and screen-analyzed, and its apparent specific gravity is determined by packing and shaking in a standard container, and weighing. The apparent specific gravity ordinarily is about 60 per cent of the true specific gravity. The unit volume present in the mill .in all tests is 700 C.C. of the packed minus 6-mesh roll product, and the number of grams occupying 700 C.C. is the unit test weight. This weight is placed in the mill dry, ground for the number of revolutions estimated to be necessary, discharged, and screened mechanically in three testing sieves of the mesh size at which the test is to be conducted. The oversize, or circulating load, is weighed, sufficient fresh minus 6-mesh feed is added to bring the total weight up to that of 700 c.c., and the charge is returned to the mill for a second grinding period. All standard ball-mill tests are conducted at 250 per cent circulating load, and the number of revolutions in the mill necessary to obtain this circulating load at any grinding period is estimated from the results of the previous period. The number of net grams of screen undersize produced per revolution of the mill approaches its final equilibrium value after several grinding periods, and this is recorded as the grindability, or relative ease of grinding, of the sample. A cylindrical ball mill, 12 by 12 in. inside, with a smooth lining and rounded corners, is used with a revolution counter Under standard grinding conditions it runs at 70 r.p.m., and contains a charge of 285 iron balls weighing 20,125 grams, ranging from I½ to ¾ in. in diameter. Tests are conducted at the mesh size to which the ore is to be ground in practice, and the capacity- to be expected from a given mill is calculated by comparing the grindability with that of a similar standard ore whose performance characteristics arc known. No satisfactory general formula for convertillg standard grindabilities into corn; rnercial power-consumption values has been

Jan 1, 1943

By A. T. Beckwith

The Lehigh Navigation Coal Company Inc. operates steep-pitch, relatively deep mines in the Panther Creek Valley, at the eastern end of the southern anthracite coal field. Commercially minable coal beds are the Mammoth,† Primrose, Orchard, and Buck Mountain;‡ their respective thicknesses average 50, 15, 10 and 12 ft. Sand slates immediately overlie and underlie these beds, though in some locations sandstone or conglomerate beds occur. Underground operations extend to a depth of 1200 ft. below the surface, and comprise a bulk of virgin work on the lower levels, and a small amount of remining on the upper levels. With increase in depth, ventilation has become a more important factor. Development In early days slopes or drifts were driven in the outcropping coal on the hillsides. Subsequent development was by water-level tunnels, with portals above high water in Panther Creek. These tunnels have been maintained because they intercept and drain much water that enters the mined-out portion of the bed and otherwise would follow down to the deep workings+ Below these tunnels, the coal was customarily developed by slopes. Muleways, which are zigzag chutes easy to travel, were also driven to connect lower openings with the water levels. When it became necessary to go still deeper, the slopes were abandoned, and vertical multicompartment shafts were sunk from the surface. Levels have been at vertical intervals of from 175 to 250 ft. Recent practice has been toward the latter spacing, splitting the total lift, however, by starting to mine from chutes extending halfway up to the level above. Chutes customarily are spaced at 50, 60 or 70-ft. centers along the gangways. At some mines, additional shafts have been sunk from the surface to the level containing the main pumping station, and these carry the discharge lines from the pumps. As the mines are ventilated by the exhausting system, both the hoisting and so-called water shafts serve as intakes. The main fan is always on the surface, and is connected with the return airways from the various levels by an air hole or a vertical shaft. For economy, when the present deepest levels were developed, and return-air shafts or air holes were not extended to the new level, but connections were made from the new level to that above by two or more holes. It is doubtful that this method will be continued, as complications in the system, increased resistance to flow, and other difficulties undoubtedly will counterbalance any economies this provision may afford. During the early years of mining, when operations were at no great depth, all development headings were driven in the coal bed. As mining progressed downward, increased ground pressures slowed develop

Jan 1, 1942

By Walter Maxson, Fred C. Bond

Since the last publication of tabulated results of grindability tests by the authors1 the total number of ball-mill tests made has more than doubled, and rod-mill tests have become increasingly important. Nearly all of the standard closed-circuit ball-mill and rod-mill tests made to date are included in the present tabulation, together with additional information not previously published. Standard Ball-mill Grindability Tests The method of conducting these closed-circuit tests has not been altered. A sample of the ore or other material is stage-crushed in rolls set at 1/16-in. opening with a 6-mesh screen until all of it has passed the screen. The combined minus 6-mesh screen undersize is mixed, sampled and screen-analyzed, and its apparent specific gravity is determined by packing and shaking in a standard container, and weighing. The apparent specific gravity ordinarily is about 60 per cent of the true specific gravity. The unit volume present in the mill .in all tests is 700 C.C. of the packed minus 6-mesh roll product, and the number of grams occupying 700 C.C. is the unit test weight. This weight is placed in the mill dry, ground for the number of revolutions estimated to be necessary, discharged, and screened mechanically in three testing sieves of the mesh size at which the test is to be conducted. The oversize, or circulating load, is weighed, sufficient fresh minus 6-mesh feed is added to bring the total weight up to that of 700 c.c., and the charge is returned to the mill for a second grinding period. All standard ball-mill tests are conducted at 250 per cent circulating load, and the number of revolutions in the mill necessary to obtain this circulating load at any grinding period is estimated from the results of the previous period. The number of net grams of screen undersize produced per revolution of the mill approaches its final equilibrium value after several grinding periods, and this is recorded as the grindability, or relative ease of grinding, of the sample. A cylindrical ball mill, 12 by 12 in. inside, with a smooth lining and rounded corners, is used with a revolution counter Under standard grinding conditions it runs at 70 r.p.m., and contains a charge of 285 iron balls weighing 20,125 grams, ranging from I½ to ¾ in. in diameter. Tests are conducted at the mesh size to which the ore is to be ground in practice, and the capacity- to be expected from a given mill is calculated by comparing the grindability with that of a similar standard ore whose performance characteristics arc known. No satisfactory general formula for convertillg standard grindabilities into corn; rnercial power-consumption values has been

Jan 1, 1943

By A. T. Beckwith

The Lehigh Navigation Coal Company Inc. operates steep-pitch, relatively deep mines in the Panther Creek Valley, at the eastern end of the southern anthracite coal field. Commercially minable coal beds are the Mammoth,† Primrose, Orchard, and Buck Mountain;‡ their respective thicknesses average 50, 15, 10 and 12 ft. Sand slates immediately overlie and underlie these beds, though in some locations sandstone or conglomerate beds occur. Underground operations extend to a depth of 1200 ft. below the surface, and comprise a bulk of virgin work on the lower levels, and a small amount of remining on the upper levels. With increase in depth, ventilation has become a more important factor. Development In early days slopes or drifts were driven in the outcropping coal on the hillsides. Subsequent development was by water-level tunnels, with portals above high water in Panther Creek. These tunnels have been maintained because they intercept and drain much water that enters the mined-out portion of the bed and otherwise would follow down to the deep workings+ Below these tunnels, the coal was customarily developed by slopes. Muleways, which are zigzag chutes easy to travel, were also driven to connect lower openings with the water levels. When it became necessary to go still deeper, the slopes were abandoned, and vertical multicompartment shafts were sunk from the surface. Levels have been at vertical intervals of from 175 to 250 ft. Recent practice has been toward the latter spacing, splitting the total lift, however, by starting to mine from chutes extending halfway up to the level above. Chutes customarily are spaced at 50, 60 or 70-ft. centers along the gangways. At some mines, additional shafts have been sunk from the surface to the level containing the main pumping station, and these carry the discharge lines from the pumps. As the mines are ventilated by the exhausting system, both the hoisting and so-called water shafts serve as intakes. The main fan is always on the surface, and is connected with the return airways from the various levels by an air hole or a vertical shaft. For economy, when the present deepest levels were developed, and return-air shafts or air holes were not extended to the new level, but connections were made from the new level to that above by two or more holes. It is doubtful that this method will be continued, as complications in the system, increased resistance to flow, and other difficulties undoubtedly will counterbalance any economies this provision may afford. During the early years of mining, when operations were at no great depth, all development headings were driven in the coal bed. As mining progressed downward, increased ground pressures slowed develop

Jan 1, 1942

By C. P. Fang, V. A. Marple, K. L. Rubow, P. H. McMurry

"It has been known for some time that sizes of aqueous particles may be affected by flow-induced relative humidity (RH) changes. The RH of air flowing through an impactor decreases as a result of pressure drop through the impactor and increases because of aerodynamic cooling within the high-speed jets, as illustrated in Figure 1. Because the sizes of aqueous particles depend on RH, the impaction stage on which such particles are collected may be different from the stage where they ought to be col¬lected, based on the particle size at the sampler inlet.Fuchs (1980) reviewed literature on water vapor condensation in impactors. The reviewed impactors operated at jet velocities of 100-250 m/s. Particle size increases ow¬ing to water vapor condensation were found in all cases. In one case (Khorguani et al., 1970), it was found that very small droplets (Dp <_ 0.02 um) containing sodium chloride and silver iodide were collected in an impactor with a cut point of 0.32 µm and a jet velocity of 100 m/s. Fuchs believed that such particles could not be captured in an impactor without having been enlarged by condensation. Winkler (1974) sampled at¬mospheric dust with a rectangular nozzle impactor that had a nozzle cross section of 0.3 x 15 mm, a cut size of 0.4 um, and an average jet velocity of 100 m/s. No adhesive was used on the impaction plate. The collection efficiency increased as air humidity was increased from 20% to 100%. The effect was especially noticeable in the increased collection efficiency for hygroscopic particles, and was attributed to condensation.Extensive theoretical and experimental research work on the distortion of the particle size distributions in impactors was done by Biswas and coworkers (1985, 1987). The impactors employed in the study were a low-pressure impactor with high jet velocity stages (Hering et al., 1978, 1979) and a conventional six-stage impactor, Sierra's model 226, operated in the incompressible regime. Ammonium sulfate aerosols were chosen for the study. In the low-pressure impactor, an orifice operating at a Mach number of 1 causes a sudden pressure drop, significantly reducing the RH in the next two stages (stages 5 and 6). This results in a 0.15-um diameter ammonium sulfate droplet shrinking to about 0.07 um at stage 5 when the inlet RH is 95%. In stage 7, the tempera¬ture drop is considerable because the jet speed is sonic. For an inlet RH of 95%, a 0.03 um droplet grows to about 0.06 um. However, the same size droplet at an inlet RH of 25% grows to only 0.04 um. For the Sierra model 226 cascade impactor that is operated with a Mach number of 0.27 at the last stage (stage VI), it was found that for an inlet RH of about 67%, particle growth would occur in stage VI. At 80% RH, both stages V and VI were affected. Biswas et al. (1987) stated that Mach number alone does not determine the size biasing; the type of aerosol, the inlet conditions, the temperature and pressure ratio, and the impactor geometry (throat length, jet-to-plate distance, etc.) must be considered for each stage in evaluating the distortion of the particle size distribution."

Mar 1, 1992