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By Baojie Zhang

The main goal of this study was not only to investigate the effectiveness of the FGX Dry Separator for removing pure rock from run-of-mine coal producing a better feed stock for conventional coal preparation plants but also to optimize the FGX Dry Separator performance for producing a salable clean coal product without using the conventional wet coal preparation processes. A Model FGX-1 Dry Separator with feed throughput capacity of 10 tph was extensively tested at the Illinois Coal Development Park using two coal samples having distinctly different cleaning characteristics. Statistically designed experimental programs were conducted to indentify critical process variables and optimize FGX Dry Separator performance by systematic adjustments of critical process variable parameters. The coal cleaning performance of the FGX Dry Separator was evaluated for the particle size range of 63.5 x 4.76 mm in most cases, although FGX Dry Separator feed consisted of nominal -63.5 mm run-of-mine coals. The best cleaning performance obtained from the FGX Dry Separator is described by specific gravity of separation (SG50) and probable error (Ep) values of 1.98 and 0.17, respectively. The SG50 and Ep values for individual size fractions were as follow: 1.90 and 0.12 for the +50.8 mm size fraction, 1.95 and 0.18 for the 50.8 x 25.4 mm size fraction, 2.01 and 0.19 for the 25.4 x 12.7 mm size fraction, and 2.03 and 0.23 for the 12.7 x 4.76 mm size fraction. These process efficiency measures produced a clean coal with ash content of 13.38% from a feed coal with ash content of 34.45%. Ash content of tailings and middlings streams were 85.09% and 39.57%, respectively. Total sulfur contents of corresponding streams were 3.87%, 4.68%, 6.87%, and 4.62%. For a relatively easy to clean Illinois No 6 seam coal (Cleaning Index: 0.72), only about 0.42% of the clean coal (i.e., 1.6 float fraction) present in the feed was lost to the tailings stream. For a relatively difficult to clean coal (Cleaning Index: 0.53), about 0.98% of the clean coal present in the feed was lost to the tailings stream.

Jan 1, 2012

By Harry E. Holzschuh

The two Agrico Chemical Company self unloading Phosphate Rock barges put into service in 1976 operate at over their design carrying capacity. Each barge now carries 26,000 short ton instead of the planned 25,000 short ton at 32 foot of draft. They are in service between Agrico's Tampa, Florida Terminal and the Donaldsonville, Louisiana Fertilizer Complex, moving a total of 2,500,000 short tons each year. This was the first use of the Parascrew design for a marine application. The barges, with their 3,000 STPH unloading and matching loading rate is the vital link in this total transportation and handling system. Rock is moved from the Pierce, Florida Mines through the Tampa Big Bend Terminal, across the Gulf of Mexico, and up the Mississippi to the Agrico Faustina Phosphoric Acid Plant 30 miles below Bacon Rouge, Louisiana. The trip from the terminal to the plant is 600 miles and the turnaround time is just about six days. Seven years operating experience has proved that the system works. Fittingly, the two barges are named the "Faustina" and the "Pierce". They have a beam of 85-feet and are 495 feet long. folded depth is 43 feet, with a ballasted light draft of 19 feet and a typical loaded draft of 32 feet. There are four separated holds with the machinery room located forward of 101 hold. Each hold is 51 feet wide at the coamings with an 18 foot straight section on the side, and a 24 foot long, 600 slope sheet for a 23 foot clear reclaim width at the bottom. The Parascrew in each of the holds is 37 feet long between the bearings and is supported on a 14 inch diameter hollow shaft. The screw pitch varies and the diameter of the flights step up from l'-9" diameter to 2'-6" diameter on the discharge end. By sloping the centerline, a minimum four inch bottom clearance is maintained across the hold. Some bending of the tube does take place.

Jan 1, 1983

By B. P. Martinez

A process for regenerating spent pickle liquor is proposed in which some of the water is evaporated at atmospheric pressure and ferrous sulfate is precipitated as monohydrate crystals..Equilibrium conditions are set for a slurry of crystals in a mother liquor containing 5 percent FeS04 and 45 percent E2S04, by weight, and boiling at about 235° F. The slurry is pumped continuously to a filter, and the mother liquor from the filter is pumped back for reuse in the pickling tanks. The net result of the process is to remove ferrous sulfate in the regeneration plant at the same rate ac it is produced in the pickling tanks. This process is integrated with the steel-pickling operation and controlled by the pickling requirements for particular steels. Spent pickle liquor is removed for regeneration continuously from both batch and continuous picklers at higher acid concentrations than would be dumped to waste because no acid is neutralized or discarded. However, the process is not designed to keep a sulfuric acid concentration stronger than 25 percent in any pickling tank. To retain flexibility in the pickling operation, all acid makeup is added to the pickling tanks. Rinse waters are used as water makeup for the pickling tanks. To further minimize waste of acid in the rinse waters, improvements of rinsing practices are outlined.

Jan 1, 1953

By N Esmaeili, M Saghaeian, L Hosseini, V Sheikhzadeh, R Hejazi

Ore roasting has long been used for beneficiation and assay upgrading of different ores. In this process, which is mainly used for sulfide ores, the ore is heated to a high temperature. More specifically, it is a metallurgical process involving gas-solid reactions at elevated temperatures. Some types of Iron ore are weak in magnetic properties. Hematite, goethite and limonite are classified in this group. Unlike magnetite, which shows strong magnetic susceptibility, these ores are hard to beneficiate through magnetic separation route. Many researches have focused on gravity separation and flotation routes for upgrading these ores and these methods have also been utilised in the industry. One of the methods to whose industrial application less attention has been paid is magnetising roasting in which the non-magnetic iron ore is transformed to magnetite which is further beneficiated through magnetic separation. In this study, magnetising roasting tests have been performed on one goethite sample in order to check the feasibility of using this method for this specific ore in industrial scale. It is worth mentioning that high intensity magnetic separation have been performed on this sample before but the achieved results were not promising. Results have revealed that through magnetising roasting magnetite content of the ore could be increased from 1.7 per cent up to about 25 per cent. FeO content of the feed sample was increased from 1.03 per cent up to around 9.25 per cent. Also optimum size distribution for achieving the best results in magnetising roasting for this sample has been determined through this study. The best results were achieved when the ore size was 0–10 mm. The effect of coal ratio, time and temperature on the process have also be studied. A coal ratio of 5 per cent, time of 30 min and a temperature of 850°C were determined to be the optimum conditions. Finally, the technical feasibility of utilising magnetising roasting for this specific ore has been confirmed.

Aug 24, 2022

By Bernardo Martins Braga

A prática de inertização de um distribuidor de 17,7 t para lingotamento contínuo de aços foi estudada numericamente através de um modelo matemático tridimensional não-isotérmico. O modelo considera uma única fase gasosa composta de dois componentes: argônio e ar. A turbulência foi levada em conta por meio do modelo k-e padrão. Os cálculos foram executados utilizando o software de CFD comercial ANSYS CFX. Diferentes configurações do sistema de inertização foram simuladas. Os resultados indicaram que a configuração atual é ineficiente e novas configurações foram sugeridas. Algumas recomendações para a correta prática de inertização também são fornecidas.

Jul 30, 2018

By A. Kelly, S. Sato

Tensile tests have been performed on aluminum single crystals of 99.99 pct purity at 196°K. The resolved shear stress when the stress-strain curve becomes concave to the strain axis depends on orientation, being always higher for crystals with an initial orientation close to the sides of the unit triangle. Observations have been made, with the optical microscope, of the appearance of slip lines on polished surfaces of the crystals. Particular attention has been paid to the first appearance of cross-slip. It is confirmed that marked cross-slip is first seen when the stress-strain curve becomes concave to the strain axis. Measurements are reported of the lengths of the slip lines observed on top and side faces of the crystals at various stages of the deformation. The lengths of the slip lines vary inversely as the flow stress for all crystals, independent of crystal orientation. From measurements of the lengths and separations of the slip traces it is possible to estimate the number of dislocations which are annihilated at each cross-slip site. It is concluded that the annihilation of screw dislocations of opposite sign is the most important cross-slip process occurring. FRIEDEL1 recently pointed out that the stress-strain curve of many pure metals of face-centered-cubic structure can be considered as composed of three stages. Plastic flow commences with a region showing a small rate of hardening which is approximately linear. This has been called "easy glide*. It is followed by a region of more rapid hardening in which the slope of the curve is again approximately constant and this, in turn, is superseded by a final part of the curve which is concave to the strain axis. Following Diehl2 we may designate these regions as stages I, II, and III respectively. The onset of stage 111 occurs at lower stresses at higher temperatures. The appearance of the slip lines formed during these various stages has been described by a number of workers?-5* The object of the present work was to make detailed observations of the lengths and separations of slip lines for a number of crystals of different orientations deformed at the same temperature and strain rate. Although much previous work has been published on the deformation of aluminum single crystals the importance of measurements of these quantities has only recently been realized—they are essential for experimental tests of recent theories of work hardening.' 9 In aluminum, deformed at room temperature, stage 11 of the stress-strain curve is very weakly marked, as the stress for the onset of stage 111 is attained soon after the end of easy glide. In the present experiments it was decided to use a temperature lower than room temperature so that stage II of the curve was better developed. However, at temperatures as low as 90°K slip lines are not measured easily with the optical microscope.5 For this reason 1960k was chosen as a suitable temperature. Even at 1960k however, stage III of the curve commences at glide strains of 6 to 10 pct and hence

Jan 1, 1960

By J. W. Cunningham, M. J. Stultzman

THE demand for improved materials for high-temperature service has increased during the last few years. Iron,- nickel,- and cobalt-base alloys have been investigated extensively and many alloys with outs'tanding high-temperature strength have been developed. Considerable research has been directed toward improved methods of processing these alloys. The vacuum furnace as a tool for the purification and improved control of high-temperature alloys has been studied extensively and production units are in operation. The research covered by this paper studied the effects of casting in air, vacuum, or argon upon the elevated-temperature properties of two nickel- and two cobalt-base alloys. The alloys investigated were Stellite 31, He 1049, Udimet 500, and Guy Alloy. MELTING AND CASTING PROCEDURES Two heats of each of the four alloys under investigation were melted and investment cast under the following conditions: 1) Stellite 31 was melted and investment cast in air and in vacuum, with an argon atmosphere for pouring. 2) He 1049 was melted and investment cast in air and in vacuum, with an argon atmosphere for pouring. 3) Udimet 500 was melted and investment cast in vacuum with and without additions of zirconium and boron. 4) Guy Alloy was melted and investment cast in vacuum and in argon. The heats were all melted in induction furnaces equipped with preformed magnesia crucibles of 50-lb capacity. The air-melting furnace was powered by a spark-gap converter of 40 kw capacity. The vacuum-melting furnace was powered by a motor-generator set of 100 kw capacity. The argon heat of Guy alloy was melted in the vacuum melting furnace under an argon atmosphere. Virgin materials were used for preparing all heats except the air heats of Stellite 31 and He 1049. The charge for the air heat of Stellite 31 was a commercial blend of shot. The charge for the air heat of He 1049 was Stellite 6 shot plus those virgin materials necessary to give the desired composition. All of the materials used were of the same grade used com-

Jan 1, 1960

By Theodore L. Turner, Joseph H. Swift

A study on a small laboratory scale was made of the use of ammoniacalammonium carbonate for batch leaching Mn from low- grade manganiferous materials ( 6 to 16 pct Mn) obtained from deposits in the southern district of Aroostook County , Maine . These materials were partially reduced at 500 ° to 700 ° C with H₂ or CO and then leached with solutions of NH and CO₂ dissolved in water . Extraction of 80 to 90 pct of the Mn was obtained from some sample materials , but 10 to 34 pct of the iron was also extracted . Material from the Littleton Ridge deposit contained manganese intimately associated with Fe , a condition which contributed to difficulties in obtaining a commercially acceptable product . Extensive investigations of the leach and reduction processes with various mixtures of H₂ and water vapor , and CO and CO₂ , failed to disclose a set of conditions for selectively leaching the manganese . Finely divided metallic Fe was probably produced during the reduction steps so that Fe was leached with the Mn in the subsequent extractions .

Jan 1, 1964

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.''~ 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

Most coal mines in Chongqing Area are highly gassy mines and liable to outburst of coal cnd methane. In order to harness mine environment and raise the Prodction capacity, methane predrainage is generally taken as a main preventive measure against outburst by all mines. In the recent ten years, the drainage volume per year has increased from 38.25 million m3 to mote than 100 million m3 with a volume over 12 m3 per tome of coal. In some mines, the draining rate is approaching to 50 percent, the rate in mining districts is up to 60 to 70 percent. The predrainage volume accounts for one fourth of total drainage volume. All these have effectively prevented the occurrence of coal and methane outbursts.This is the current situation of methane predrainage in coal mines in Chongqing area, the methods of predrainage to prevent outbursts and related parametres comprise:1. methods, technologies, supervision and effect of predrainage in different geological and mining conditions,2. the change of methane parametres pre-and post-drainage and the analysis of the effect of methane predrainage to prevent outbursts, and3. ways to improve the effect of methane predrainage.Thus the safety, reliability and production of Changqing Coal Mines will be enhanced.

Jan 1, 1991

By K. M. Deen, E. Asselin

Hydrometallurgical processes, i.e., Cu extraction from a mineral concentrate and Zn electrowinning, are coupled in a battery-like system. A naturally-sourced CuFeS2 concentrate mixed with activated carbon is used as a positive slurry electrode separated by a membrane from the negative compartment. The Zn is deposited on the negative Al current collector from the circulating solution containing 100 g L–1 Zn2+ in 0.2 M H2SO4. In this battery-like system, both CuFeS2 oxidation and Zn deposition are possible during the charging cycle. The additional benefit of this setup is the energy storage if discharged by re-dissolving the Zn in the negative compartment. Cyclic charge/discharge testing revealed oxidation of the mineral concentrate supported by the Zn2+/Zn redox reaction. The coulombic and energy efficiencies increased monotonically to 95.6% and 43%, respectively, with approximately 7% Cu extraction during the first 10 charge/discharge cycles (completed in 2.4 hours). The discharge energy density also increased from 2.7 to 36.2 mWh L–1. In the following 90 cycles, the low energy (~14%) and coulombic (~33%) efficiencies that were obtained confirmed the progress of irreversible reactions e.g. Cu extraction from the mineral concentrate. Overall, approximately 16% Cu was extracted from the mineral concentrate over 100 charge/discharge cycles (or about 12.2 hours). INTRODUCTION Chalcopyrite (CuFeS2) is the main economic copper-bearing mineral, which is generally processed through pyro-metallurgical methods to produce Cu. However, environmental challenges with smelting, depletion of high-grade ores and a large number of impurities (some toxic) have resulted in continued research aimed at the development of hydrometallurgical technologies to process CuFeS2. There are a number of different proposed hydrometallurgical routes in the literature to treat CuFeS2 concentrates, but only a few have attained commercial acceptance (Dreisinger, 2006). Zn is another commercially important metal that is mainly produced via acid leaching of roasted sphalerite (ZnS). The pregnant leach solution is rigorously purified to remove all the soluble metal impurities, particularly those that are electrochemically more noble than Zn and present large kinetic activity towards H2 evolution (Hu & Piron, 1992; Mackinnon, Brannen, & Fenn, 1987; Mureşan, Maurin, Oniciu, & Gaga, 1996). Metal impurities such as Cu, Co, Ni, Cd, Pb, Sb, As, Ge and Fe etc., if they exist in the leach liquor beyond the tolerance limits, can decrease the current efficiency during electrowinning and deteriorate the deposit quality (Alfantazi & Dreisinger, 2001). A complete flow sheet, describing the process steps of Zn metal production from mine site to pure metal is presented elsewhere (Brown, Meisenhelder, & Yao, 1983).

Jan 1, 2019

By Omar A. Muhtadi

Brohm Mining Corporation, in conjunction with Geobiotics, Inc. has nearly completed a pilot heap biooxidation test on sulfidic ore from the Gilt Edge Mine near Rapid City, South Dakota. Some results from the pilot testing have shown the correlation between the extent of biooxidation and the increase in cyanide gold recovery seen in the laboratory. Bench scale testing showed approximately a 1: 1 correspondence between percent biooxidation and percentage increase in gold recovery. The economics of large scale biooxidation were also evaluated in terms of increased gold recovery versus the time and cost of biooxidation. In the pilot test, the rate of biooxidation became limited by the acid water neutralization system. Later in the test, measures were taken to overcome this problem. Results from bench scale column bioleach testing showed that the Gilt Edge sulfide ore is quite amenable to biooxidation with a substantial increase in gold recovery with just partial biooxidation. Direct cyanide gold recovery of 51 % on ore crushed to -3/8" was enhanced to 82-85% after oxidation of37% of the sulfide as measured by iron. Laboratory experiments also showed that the slow rate of biooxidation of this ore could be accelerated by removal of metal sulfate build up from the leach liquor. The combination of these normally non-toxic metals became inhibitory to biooxidation if not removed by the neutralization circuit.

Jan 1, 1995

By James Hance

INTRODUCTION SEVERAL years ago the writer was connected with the Mint and Assay Service of the Federal Government as Assistant Assayer at the-Salt Lake Assay Office. At that time cyanide bars formed approximately half of the bullion purchased by that office. Disagreement in valuation between the producer and the office was not infrequent and to a lesser extent between the Assay Office and the Mint, this latter issue being soon obviated, however, by adopting uniform methods of sampling the bars. A nice problem seemed to offer itself for research and the writer began a series of experiments which were soon terminated, however, by his going into other work, and which he regrets have not since been completed. The results attained may suggest lines of approach to this problem and are, therefore, offered in this article. Frederick P. Dewey, Assayer of the Mint Service, U. S. Treasury, has had abundant opportunity to study this phase of metallurgy, and has published several articles relative thereto.1 In the last article cited Dewey quotes analyses of cyanide bullion similar in content to that with which the writer worked, and which show some of the variations met with. In the discussion of these papers,2 comparison has been made in terms of copper bullion, but to the writer this seems of questionable value. Copper bullion may resemble silver bullion, but its similarity to, gold bullion with little or no silver content is another thing. The freezing-point diagram of Roozeboom, reproduced on page 881 of the discussion cited, is for the gold-silver series as stated. The original problem under discussion, however, is not this alloy but gold with base metals other than copper. As for the statement that the corners are the worst places or

Jan 2, 1916

By John K. Alley, Robert C. Johnson

The primary objective of the research was to grow aluminum borate whiskers or fibers and to determine their properties . Molten salt bath and vapor deposition methods were used for synthesis . Whiskers exceeding 25 mm in length were grown . The best results were obtained by vapor deposition and most of the data presented were determined on whiskers produced by this method of synthesis . Three different aluminum borate compositions ( 4.8A1203 B₂03 , 3.8A1203 B203 , and 2.5A120, B203 ) were identified by X- ray and chemical analyses . The aluminum borates 4.8A1203 B20, X- ray diffraction pattern was indexed . Optical data were obtained on the three aluminum borates . In addition the following whisker properties and growth factors were determined : ideal atmospheric dewpoint for vapor deposition method , near 65 ° F ; temperature range of whisker formation , 1,000 ° to 1,400 ° C for the molten bath method of synthesis , > 1,000 ° to < 1,400 ° C for the vapor deposition method ; whisker growth rates by the vapor deposition method , maximum 12 mm per hour ; raw material vaporization , 50 to 60 percent of the fluoride lost in 1 hour ; tensile strength , maximum obtained 174,162 psi ; melting points , ranged from 1,675 ° to 1,795 ° C.

Jan 1, 1965

By Janet Torma-Krajewski

Job tasks that involve exposure to work-related musculoskeletal disorder (WMSD) risk factors may impact both the risk of injury and production downtime. Common WMSD risks factors associated with mining tasks include forceful exertions, awkward postures, repetitive motion, jolting and jarring, forceful gripping, contact stress, and whole body and segmental vibration. Mining environments that expose workers to temperature/humidity extremes, windy conditions, and slippery and uneven walking surfaces also contribute to injury risk. National Institute for Occupational Safety and Health (NIOSH) researchers worked with powder crew members from the Bridger Coal Company to identify and rank routine work tasks based on perceived exposure to WMSD risk factors. This article presents the process followed to identify tasks that workers believed involved the greatest exposure to risk factors and discusses risk reduction strategies. Specifically, the proposed prill truck design changes addressed cab ingress/egress, loading blast holes, and access to the upper deck of the prill truck.

Jan 1, 2009

By Ping Zhang, Xu Ma

Perturbations arising from mining operations significantly affect the stability of rock masses, and the influences aggerates with the rapid increase of mining-operation depths during recent years. The subsurface structures with major discontinuities subject to seismic hazards resulted from the shear-slip behaviors of rock masses. In order to identify the shear-slip regime of discontinuities and calculate seismic moment and seismic energy involved with shear-slip behaviors, we use discrete element modeling to study the shear slip failure along discontinuities in an underground mine. The recorded characteristic and properties of sub-contacts in DEM provide a basis for computing and visualizing the temporal and spatial distribution of seismic moment and seismic energy with mining operations. We computed the seismic energy and seismic moment using the numerical modeling method and the analytic method. We compared the result of summing seismic energy and seismic moment from the subcontacts of numerical models and the result of the analytic method. We confirmed that this tool can be used in comparative analyses. We also found that seismic moment and seismic energy, associated with shear stress drop and shear displacement increase, accumulate in the vicinity of major discontinuities. Mining operations at a greater depth cause greater changes of seismic moment and seismic energy, leading to a higher risk of inducing seismic hazards. Quantifying seismic potential using discrete element modeling can greatly facilitate the investigation of instability of geological discontinuities and thereby can help estimate the potential of seismic hazards.

Jan 21, 2023

By Martin Hertzberg

A comprehensive study of the flammability behavior of air-dispersed coal dust was made using an 8-liter Bureau of Mines system which included the following: (1) An optical probe to continuously monitor the dust concentration; (2) measurements of ignition energy requirements; (3) the use of several narrow size distributions ranging from 3 to 65 µm in average diameter; (4) measurements of the extent of flame propagation based on pressure attained at constant volume, residual O2 content after combustion, and flame and dust surface temperatures. The measured lean limit for near-horizontal propagation of Pittsburgh coal dust was about 135 mg/l (milligrams per liter), virtually independent of particle size. Required ignition energies were hundreds of joules and measured limit flame temperatures were 1,450 to 1,600° K. Data on inerting? requirements with KHC03, rock dust, and NaC1 gave approximately the same results as full-scale explosion tests. The good agreement suggests that the cause of earlier discrepancies between small-scale laboratory data and full-scale mine experiments may not have been the difference in size, but were rather a problem of proper care in isolating and controlling the significant experimental variables and in consistently defining the extent of flame propagation at the limits.

Jan 1, 1979

The movements of the major continental cratons and the distribution of the principal tectonostratigraphic terranes are reviewed in the Circum-Pacific region during Phanerozoic time. It is concluded that major periods of orogeny and consolidation of accreted terranes in the Circum-Pacific region is best correlated with periods of advance of cratonic blocks over the Pacific Ocean basin combined with relative convergent to transpressive motions between the cratons and the Pacific Ocean plates. Continental collisions and collision of terranes with continental margins has not been a significant geodynamic process in Circum-Pacific tectonic evolution. The permanency of the Pacific Ocean basin throughout 500 MA is remarkable. It appears no continental block or large continental fragment has ever entered or crossed the Pacific Ocean basin since early Paleozoic time. This area where no continent has entered for 500 MA is almost coincident with the present Pacific Ocean basin and seems associated with the 25 to 100 meter geoid high which dominates the present-day figure of the Earth.

Jan 1, 1990

This study is to compare reinforcement of rebar and spiral bolts using laboratory pull-out test. For this purpose, the pull-out test was performed in laboratory for a curing period of 7- and 28-days. Based on the results of the laboratory pull-out test, the pull-out load of two support members rapidly increased with displacement until the breakpoint and then gently increased with displacement. The maximum pull-out load of spiral bolt was larger than that of rebar bolt. This implies that a spiral bolt is more stable than a rebar bolt for securing the stability of a ground or rock mass when a support was installed in a ground or rock mass under the same condition. In addition, the maximum external pressure of two support members increased with pull-out load and with displacement and the distribution of external pressure was commonly larger at the upper part than at the lower part of specimen. This indicates that the bond failure location by pull-out load should be mainly occurred at the surface portion of specimen. In conclusion, the reinforcing effect of a support depends on a curing period and a type of support member, and reinforcement of a spiral bolt is more effective than that of a rebar bolt.

Dec 6, 2010

By Isaac A. Zlochower, Michael J. Sapko, Ronald S. Conti

The Bureau of Mines has developed a system using a high-speed electropneumatic mechanism for the rapid (grab) sampling of dusts and gases during an explosion. The sampling system consists of an aluminum housing that incorporates two 30-cm3 preevacuated glass vials with rubber septums. Upon actuation, the sampling probe needle is driven through the septum with a pressurized air pulse, filling the tube with gas and dust from the mine explosion. After a predetermined time, the sampling probe needle is retracted by a second high-pressure air pulse to its normal (quiescent) state, allowing the sampling tube to reseal. The onset and the duration of sampling are independently variable and controlled by a time-delay relay package. This technique enables the monitoring of pyrolysis-charring in coal particles and the generation and combustion of the pyrolysis vapors in both large- and small- scale explosions. The following results were obtained from full-scale dust explosion tests at the Bureau of Mines Lake Lynn Laboratory: (1) Gas sampling of the leading edge of the flame front shows the large concentration changes, characteristic of the flame front; (2) gas samples taken entirely in the flame zone consist of pyrolysis and combustion products with very low residual oxygen; and (3) the particles collected in the flame zone show signs of extensive pyrolysis and charring.

Jan 1, 1988