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By Sidney L. Hansen

This report was written because of concerns of the possible error involved when stopping leakage is incorrectly assumed to be turbulent. It is common practice to use turbulent flow to characterize each airway branch for computer models of mine ventilation circuits. This report addresses computer methods to accurately model air-leakage paths. A graphical method is described for deriving the equations of leakage through stoppings. The significance of modeling leakage according to the proper flow regime (turbulent or laminar) is demonstrated by the following method. A hypothetical mine was developed to represent present-day conditions. The stoppings of the mine were chosen to leak according to laminar flow properties. The mine was represented by constructing two base models. One base model used turbulent leakage equations for each stopping, while allowing the corresponding resistances to vary. The other base model used the correct laminar leakage equations and held corresponding resistances constant. The mine circuit was then expanded to demonstrate future development. The expanded portion was designed with stoppings that were identical to those in the base mine. The mine in the future state was represented by two projection models. Each projection model was an extended version of its associated base model; one used turbulent leakage equations and one used laminar leakage equations for stoppings. The laminar projection was theoretically correct and is used as a basis for comparison of differences in airflow quantities. Rather than modeling every leakage path for stoppings, a more practical approach is also discussed. A good level of accuracy is maintained by combining leakage paths if the proper flow equations are used. This method was used on the hypothetical mine projection and the flows were very similar to the theoretical values.

Jan 1, 1989

The U.S. Bureau of Mines investigated CaF2-assisted H2SO4 leaching to extract Mn from domestic silicate manganese resources. A pyroxmangite ((Mn,Fe)Sio3) ore containing, in pct, 34 Mn, 23 Si, 3 Ca, and 1 Fe was leached at 90° C for 3 h in a solution containing, in g/L, 12 Mn, 140 H2SO4, 130 (NH4)2SO4, and 5 CaFz. Manganese extraction of 92 pct was achieved. Manganese of greater than 99 pct purity was prepared by electrolyzing the purified pregnant solution at a current density of 50 mA/cm2. The current efficiency was 65 pct.

Jan 1, 1991

By M. L. F. Gameiro

The aim of this study was to check the feasibility of non-dispersive solvent extraction to recover copper from ammoniacal solutions using hollow fibres as contactors. Data of mass transfer using the B3-diketone LIX 54 as extractant was obtained. The influence of tube and shell hydrodynamics on the extraction rate of copper was studied. The overall mass transfer coefficient of copper was calculated from the experimental data by performing steady state material balance. Simultaneous extraction and stripping experiments were carried out using different volume ratios of feed phase to stripping phase. The results obtained showed that practically all the copper content was removed from the ammoniacal feed solutions. The recovery of copper reached nearly 100% in several extraction-stripping experiments.

Jan 1, 2007

By Carlos A. Nunes

V-AI alloys have been used commercially to produce pure vanadium by Electron Beam Melting (EBM), as well as to obtain Ti-based alloys (ex. Ti-6AI-4V). The aluminothermic reduction of vanadium pentoxide (V205) is the primary method to obtain these alloys. Concerning the obtention of pure vanadium, earlier studies have indicated alloys presenting Al content between 10 and 15wt% as the most suitable alloys for direct EBM. In the present work we investigated the aluminothermic reduction of V205 in order to define process parameters for the obtention of V-AI alloys with composition in the previously mentioned range. It was studied basically the effect of excess AI, CaO addition and reaction atmosphere (air, argon) on the alloy composition (AI, 0, N). In addition, microstructural analysis of selected alloys by SEM (BSE images) and EPMA (WDX) were also performed.

Jan 1, 1996

This paper gives an overview of the difficulties associated with determining the strength of hard rock pillars. Although a number of pillar design tools are available, pillar collapses still occur. Recent examples of large scale pillar collapses in South Africa were caused by weak partings which traversed the pillars. Currently two different methods to determine the strength of pillars are used namely, empirical equations derived from the back analyses of failed and stable cases and numerical modelling tools with an appropriate failure criteria. Both techniques have their limitations and additional work is required to obtain a better understanding of pillar strength. Empirical methods are popular and easy to use, but care should be exercise that the results are not extrapolated beyond the range of the data which were used to derive them. An example is the Hedley and Grant formula (derived for the Canadian uranium mines) that has been used for many years in the South African platinum and chrome mines (albeit with some adaptation of the K-value) to design the pillar layouts. Very few collapses have nevertheless been reported and it might even be too conservative an assumption of pillar strength in these mines. As an alternative, some engineers strongly advocate the use of numerical techniques to determine pillar strength. A close examination unfortunately reveals that these techniques also rely on many flawed assumptions. An area where numerical modelling is invaluable, however, is to study specific pillar failure mechanisms, such as the influence of weak partings on pillar strength. In conclusion, it appears that neither empirical techniques nor numerical modelling currently provide a solid basis to conduct pillar design. It is therefore recommended that both these techniques be utilised when confronted with a pillar design to obtain the best possible insight into the problem.

Nov 25, 2010

By F. R. Archibald

IN a recent paper on Roasting of Beattie Concentrate (1), the importance of controlling the temperature, and the sulphur dioxide concentration was pointed out. It is the purpose of the present paper to present further information by which the most satisfactory conditions for roasting are more specifically defined. Incidentally, a method of illustration is developed for representing roasting curves on a three-dimensional diagram. The primary purpose of this investigation, as in the case of the earlier work, was the search for information that would be of practical value in improving plant operation and gold recovery. For this reason, emphasis is laid on the regulation of essential variables and on definition of their optimum values and limits for maximum gold recovery in the subsequent cyanidation process. Good roasting is considered at least as much an art as it is a science, and, in this light, theoretical consideration of the nature and sequence of the chemical reactions involved is of secondary importance. The presentation which follows includes a series of graphical illustrations, an interpretation of the graphs, some theoretical considerations, and a description of laboratory technique.

Jan 1, 1940

By Mike Dry, Mani Manimaran, Mykolas Gladkovas, Mariam Melashvili, Chris Fleming

Development of a new hydrometallurgical flowsheet requires quantitative description and optimisation of all chemical processes in each step in the process, as well as mass and energy balances across a fully integrated circuit incorporating all the unit operations. This paper covers steps leading to early modelling of a gold processing plant based on a bromine/bromide lixiviant. Research conducted at SGS Lakefield in collaboration with Albemarle provided the data that were used in the development of this model. A previous investigation in this series described the leaching behaviour of various mineral types in Stabilized Bromine reagent. This paper is concerned with the gold recovery operations associated with a conceptual bromine leaching process. The recovery of gold by different types of resin and activated carbon was studied, along with various methods of stripping the gold from the extractant. The circuit is closed by electrolysing the spent leach liquor to oxidize some of the contained bromide back to bromine that would allow recycling the regenerated leach solution back to the leach step in the process.

Jan 1, 2015

The Bureau of Mines found sulfur hexafluoride (SF6), released from a lecture bottle, to be an ideal gaseous tracer for studying mine ventilation systems. Air samples were collected in glass syringes and analyzed by electron-capture gas chromatography. In these studies, the lower limit of detection was about I part per billion by volume (ppb) of SF6 per part of air. Experiments conducted in a Pennsylvania limestone mine showed this technique useful in evaluating the effectiveness of auxiliary fans, measuring low flow velocities, probing the air circulating near a working face region where ventilation appeared to be poor, and estimating volumetric flow rates in airways of large cross-sectional area and having low flow velocities. This technique was also used in a western vein-type metal mine to measure the amount of return air being recirculated into the intake air due to leakage through old stoped areas.

Jan 1, 1974

By Henry N. Toler

Activity in the search for oil and gas in Mississippi was greater than it had been for several years. Interest manifested and exploration work in wildcat areas were at an all-time peak. Drilling activity in the state's two proven fields continued to decline; there being only four wells drilled in the Jackson field during the year. At the end of 1937 it seems very likely that 1938 will be a big year in the search for oil in Mississippi. Deilling Activity Jackson Gas Field.—The four wells drilled in the Jackson field during 1937 resulted in three producers and one dry hole. Nine producing gas wells were abandoned during the year because of salt-water encroachment. The total production for 1937 was 14,248,000,000 cu. ft., which was a billion and a quarter increase over 1936, the highest year prior to 1937. The grand total produced from the field is 72,341,000,000 cu. ft. of gas and over 15,000 bbl. of oil. The three wells that had produced oil the past few years were also abandoned during the year because of the small quantity of oil and the large volume of salt water.

Jan 1, 1938

Broken Hill has a long history of mining and it is without doubt a world-class deposit. It has seen production since 1885, notwithstanding the strikes that made the city infamous in days gone by and has given rise to three of the nation's leading mining companies; BHP, North and CRA. Prior to the formation of Pasminco in 1988, the mining and exploration activities were undertaken by separate companies who were business competitors and tended to operate in relative isolation. There were a few notable exceptions such as the Central Geological Survey of 1936 - 38 (Gustafson, 1939; Gustafson et al, 1950) and the Mine Managers' Association. The latter sponsored many university based research projects. At a local level there was good communication and cross fertilisation within this remote outback centre through personal contacts and professional bodies such as The AusIMM. Pasminco was formed by the merging of the lead-zinc interests of CRA Ltd and North Broken Hill Ltd. In Broken Hill this meant the joining together of North Mine and ZC Mines, the two largest mines then operating on the Line of Lode, to form Pasminco Mining Broken Hill Operations. The exception was the central part of the orebody, the original BHP, South Mine and

Jan 1, 1997

By Maria I. De Rosa

For all coal mining categories, data on coal mine fires during 1990-1999 have been reported as actual numbers and calculated values. 1. For each mining category, actual numbers include the total number of fires, fire injuries, employees' working hours, lost workdays, and coal production (for underground and surface mines only) for a 1 0-year period (1 990- 1999) and for five successive 2-year periods with the 10-year period. These numbers have also been reported by state ( 10-year period). The actual number of fire fatalities has been reported by time period. Furthermore, actual numbers of fires for the five 2-year periods have been reported by ignition source, method of detection and suppression, equipment involved, location, and burning material. Actual numbers of fire injuries per number of fires causing injuries and total fires have been reported by year, ignition source, equipment involved, and location. 2. For each mining category, the calculated values include the fire and injury risk rates during the 10-year period and the five 2-year periods. The fire risk rate (Frr) values were calculated according to the USBM formula [Pornroy and Carigiet 19951. The injury risk rate (Irr) values were calculated according to the MSHA formula [MSHA 1991a, 1992, 1993, 1994a, 1995a, 1996, 1997, 1998b, 1999c, 20001. Also, risk rate values for individual states (10-year period) were calculated according to the above-mentioned formulas. Of note is that only the risk rate values for the 10-year and five 2-year periods and risk rate values for individual states with the highest number of fires and fue injuries were considered for comparison purposes. The fatality risk rate values were not calculated because of the extremely small number of fire fatalities during the 1 0-year period. 3. Calculations of risk rate values are as follows: a. Fire risk rate (Frr) value: Number of fires per million tons of coal produced [Pornroy and Carigiet 19951. b. Injury risk rate (Irr) value: Number of fire injuries multiplied by 200,000 working hours per total employees' working hours [MSHA 1991a, l992,1993,1994a, 1995a, 1996, 1997, 1998b, 1999c, 20001. The Irr value is the average risk rate value for the number of fire injuries per 200,000 working hours for a given time period.

Jan 1, 2004

By Lloyd M. English

Accurate knowledge of the gas content of coal seams has always been important, according to the work of W. P. Diamond and J. R. Levine, 1981, but is becoming increasingly important for several reasons: Excessive gas emissions in the mine can create explosion hazards. Depletion of shallower coal reserves is forcing underground operations to greater depths where coal gas content is often greater. Natural gas prices and the need for new energy sources make naturally occurring coal gas a potential resource of economic interest. Information on coal gas content will impact mine designs, mine ventilation system specifications, and the effectiveness of methane control procedures used. Accurate resource assessment is needed for projecting and designing potential gas utilization enterprises. The emphasis upon methane hazards in mining operations, and upon potential methane utilization, has led us to improve upon the procedures used to determine coalbed methane levels.

Jan 1, 1984

By D. Harry Chambers

The Bureau of Mines, through a memorandum of agreement with the Department of Defense, conducted research to recover silver from aircraft scrap. Silver was recovered by an electrolytic method from stainless steel honeycomb sections separated from the aircraft. These sections had been constructed by sandwiching a stainless steel honeycomb core between stainless steel sheets and then brazing the assembly with a silver alloy. Over 300 pounds of silver was used in the B-58 bomber, concentrated in certain honeycomb sections of the aircraft. Following shredding of the aircraft parts, an average of 95 percent (ranging 67 to 100 percent) of the silver was recovered in a single electro-refining step. After the electrorefined product was magnetically cleaned, purity of the recovered silver was greater than 99.3 percent.

Jan 1, 1980

By Lénaick Menot, Anne-Sophie Alix, Charline Guérin, Sébastien Ybert, Ewan Pelleter, Florian Besson

Ifremer, on behalf of the French government, holds an Exploration Contract with the ISA for polymetallic nodules in the CCFZ. As part of its exploration activities, the NODULE (2015) cruise was conducted to acquire a 50 m resolution bathymetric coverage of the eastern sectors of the Contract. In 2017, Ifremer initiated an update of the Contract’s mineral resource estimate. Metal grades and nodule abundance were modelled using ordinary kriging. The inferred mineral resource is 436 Mt (wet) of polymetallic nodules at an average abundance of 10.7 kg/m² with a cut-off grade of 7 kg/m². Based on the geological data, seamounts, sedimentary basins and steep slopes contain no nodules. Technical studies of most of the contractors have determined that nodule collectors could not maneuver on slopes above 5 to 10°. Thus, Ifremer has selected a quite conservative slope threshold of 7 % (about 4°) and a backscatter threshold of -29.5 dB to delineate nodule-free areas. This has a significant impact on the resource estimate, as about 54 % of the eastern Contract areas are excluded. Near-seafloor high-resolution mapping and photographic transects, associated with finer-spaced box-core sampling could constitute a next step to better constrain the mineable areas and to get a higher level of confidence in the resource estimation.

Jan 1, 2018

By R. Karl Zipf

Multiple seam mining interactions caused by full extraction mining, whether due to undermining or overmining, frequently involve tensile failure of the affected mine roof. The adverse ground control conditions may prevent mining for both safety and economic reasons. Prior researchers have identified the geometric, geologic and mining factors controlling multiple seam mining interactions. This numerical study examines the mechanics of these interactions using a modeling procedure that 1) incorporates the essential constitutive behavior of the rock such as strain-softening of the intact rock and shear and tensile failure along bedding planes and 2) captures the geologic variability of the rock especially the layering of weak and strong rocks and weak bedding planes. Specifically, the numerical study considered the effect of vertical stress, interburden thickness, and the immediate roof quality of the affected seam in both undermining and overmining situations. The models show that for overburden-to-interburden thickness (OB/IB) ratios of less than 5, interactions do not occur, and that for OB/IB more than 50, extreme interaction is a certainty. In between, the possibility of an interaction was found to depend on gob width-to-interburden thickness ratio, site specific geology and horizontal stress to rock strength ratio in addition to the OB/IB ratio. The models also showed that horizontal stress was profoundly altered well above or below a full extraction area and that these changes are likely to influence the success or failure of multiple seam mining. The role of horizontal stress in multiple seam mining interactions has received little attention in prior investigations. Four factors control the mechanics of multiple seam mining interactions: 1. vertical stress concentration 2. horizontal stress concentration 3. stress re-direction 4. bedding plane slip bands A combination of vertical and horizontal stress increase and high stress gradients in the vicinity of full extraction areas re-orients principal stresses into a very unfavorable direction. This seemingly small stress re-orientation has a profound adverse effect on bedded rock.

By L. Dorador

"In block caving projects, the rock mass fragmentation process plays a controlling role in the design and success of the operation. Poorly understood, however, is the secondary fragmentation that occurs as blocks of caved ore move through the draw column and interact with one another. The initial size of these blocks can be considerable, as can their strength and stiffness as the presence of non-persistent joints and veining will be variable. Intuitively, more planes of weakness within the blocks should facilitate early fragmentation, but this assumption could depend on additional factors such as contact type between adjacent blocks and loads imposed through varying heights of the draw column. This paper reports the findings from a laboratory investigation examining the role of block strength and planes of weakness on secondary fragmentation. Testing was carried out on small-scale concrete cubes of uniform size, with and without embedded planes of weakness. The results obtained demonstrate the significance of veining and other planes of weakness on secondary fragmentation under one-dimensional compression, and allow for the development and constraint of empirical guidelines.INTRODUCTIONThe fragmentation process during block caving involves the initial (primary) fragmentation of rock in the cave back through stress-induced fracturing, release and fall of individual blocks onto the broken ore muck pile. This is followed by secondary fragmentation involving splitting and rounding of blocks as they move downwards through the draw column towards the draw point. The degree of secondary fragmentation achieved involves many variables related to the mechanical properties of the rock blocks as well as different operational factors. With respect to the block properties, the presence of planes of weakness in the form of veining and non-persistent joints is a topic of special interest given the significant influence it potentially has on the overall fragmentation process. However, the inherent heterogeneity and variable size of individual blocks generated from the cave back makes it difficult to develop any empirical relationships relating secondary fragmentation to compressive loading based on insitu observations and mine data. More preferable would be to keep constant block parameters, such as shape (aspect ratio and angularity) and size, through controlled laboratory testing. This work presents an efficient technique to fabricate concrete cubes in order to keep constant different block properties. These are then tested under one-dimensional compressive loading. The objective of these tests is to emulate a tight packing condition of single blocks such as that which would occur when the air gap between the cave back and broken ore in the draw column is small. When the air gap is small, detached blocks are unable to fall and rotate but instead maintain a very tidy arrangement with minimal voids left between adjacent blocks."

Jan 1, 2015

By T. F. Witherbee

Success in this operation is much promoted by adhering to a definite plan. Random work is of little account. In general, anything done that will enable the blast to pass through the obstruction, providing the melted products can be got away from the tuyeres, is all that is necessary. Hence the utility of high explosives in breaking up the chilled mass in the hearth. (This, by the way, should be done with the wind on, so as to prevent the filling of the tuyeres with the usual rush of cinder following the explosion.) If from any cause it is not expedient to use explosives, the simple but very effective instrument known as the kerosene or hydro-carbon blow-pipe can be depended upon to perforate the contents of a blastfurnace hearth to any extent required. It consists, as used in the Champlain district (first by Mr. Charles H. Foote), of several pieces of gas-pipe, liberally supplied with elbows (five in number) so as to be flexible; one end being connected to a tuyere-stock and the other brought to bear upon the place to be melted out. A 1/4-inch pipe leads from the last length, or blow-pipe proper, to a reservoir of kerosene or crude oil, set high enough to overcome the blast-pressure, This 1/4-inch pipe should have a valve conveniently located to regulate the flow of oil, and also another next to the reservoir. With a good hot blast and a surprisingly small quantity of oil, this apparatus will melt, burn or mechanically dislodge anything around a blast-furnace hearth excepting firebrick and water-cooled castings or plates. It may be helped by vigorous prodding with a suitable steel bar, to detach the fused but sometimes sticky material. A very good way to operate on a chilled hearth with it, is to open up a hole from a cinder- or iron-notch to the nose of the most convenient tuyere and then make a blow-pipe of the tuyere by turning

Jan 1, 1885

By M. M. Fine

Research on the solid-fuel reduction of Knob Lake iron or was conducted to upgrade lean and high-grade ores and recover metallized products suitable for ironmaking or steelmaking. The program included bench-scale tests plus a number of continuous campaigns in a 34-inch by 36-foot rotary kiln. In the laboratory, metallizations in excess of 90 percent were achieved by reduction of crushed ore with lignite at 950° C. Subsequent grinding (to minus 200 mesh) and wet magnetic separation yielded concentrates analyzing less than 3.0 percent SiO2 and more than 90 percent Fe. Virtually complete recovery of the iron occurred. During early continuous tests the crushed ore created a serious ringing in the kiln which was solved by balling the fines with bentonite. Unfortunately, at this temperature (950° C) the bentonite militates against the concentratability of the metallized product. For example, a run on direct-shipping ore produced concentrates of 89.5 percent Fe and 4.5 percent SiO2. At 500° C the adverse effect of bentonite disappeared and concentrates of 88.4 Fe and only 2.5 percent SiO2 resulted. However, with comparable feed r Los, the metallization dropped to 73 percent.

Jan 1, 1969

Albin, B. R. Billings, Mont. '28 Amster, Nathan L 25 Broad St., New York, N. Y. '28 Bache, Jules S 42 Broadway, New York, N. Y. '28 Barrett, C. P 5338 Harper Ave., Chicago, Ill. '28 Baum, Frank G., Cons. Hydro-Electric Engr Cassel, Shasta Co., Cal. '28 Bayless, William S Russ Bldg., San Francisco, Cal. '28 Bean, Louis H., V.P., Dwight P Robinson & Co., Inc. 125 E. 46th St., New York, N. Y. '28 Beatty, Ross J 1105 First National Bank Bldg., Chicago, Ill. '28 Benwell, Oswald F 110 Gilpin St., Denver, Colo. '28 Beymer, Albert S., V P., Keystone National Bank Pittsburgh, Pa. '28 Blumenthal, Hugo 44 Pine St., New York, N. Y. '28 Boettcher, Charles Ideal Cement Bldg., Denver, Colo. '28 Bowlus, Thomas H Iola, Kans. '28 Brown, Colvin B. .. Chamber of Commerce, Washington, D. C. '28 Burt, Maxwell Struthers Southern Pines, N. C. '28

Jan 1, 1923

By L. S. Austin

I. Introduction............ 431 II. Organization............ 432 III. Production............ 434 IV. Transportation............ 435 V. Sampling..........436 VI. Concentration............ 440 VII. Blast-Furnace Plant........... 442 1. Construction........... 442 2. Operation............ 450 VIII. Advantages of Large Furnaces........ 455 1. Saving in Fuel........... 455 2. Saving in Jacket.Water..... 455 3. Quick and Large Discharge of Matte and Slag... 455 4. Decrease of Incrustation... 456 5. Elasticity of Operation..457 6. Large Flow Through Fore.Hearths...4. 57 7. Alterations Without Interruption..458 8. Variations in the Composition of Slag....458 9. Less Labor-Cost.......... 458 10. Less Initial Cost.......... 460 IX. Briquetting.Plant........... 460 X. Roasting-Plant............ 462 1. Kilns............ 462 2. Reverberatory Furnaces...468 XI. Converter-Plant........... 474 1. Construction ..........474 2. Operation.......... 476 XII. Flues and Stacks........... 478 XIII. Arsenic-Plant............ 480 XIV. Coke-Washing Plant.... 452 XV. Slime.Ponds ........... 483 XVI. Brick-Plant............ 484 I. Introduction. The Washoe plant; in Anaconda. Mont., together with the local street-railroad, ranches. a foundry and machine-shop. A

Jan 1, 1907