Search Documents

By William C. Cooley

This report includes a review of all relevant data on underground hydraulic coal mining technology collected during Phases I and II of the contract. Data collection and analysis procedures are discussed. Information is included on past and existing underground hydraulic coal nines in the USSR (27 mines), Poland (1 mina), Germany (1 mine), Japan (1 mine), Canada (1 mine), China (1 nine), as well as underground tests conducted in the United Kingdom (2 mines), and in the United States (5 mines). Data is also included on Czechoslovakia and New Zealand. In all operational hydraulic mines, coal is transported underground in flumes. In most cases, bituminous coal is broken by monitors with nozzle diameters larger than 16 mm (0.63 in.) and jet pressures of 50 to 125 atm (710 to 1780 psi). Technical design and performance data are included on all key types of equipment and components which have been developed for, used in, or which could be potentially be used in hydraulic coal mining systems. Data sheets are presented on 180 types of equipment, divided into 24 major categories. Where available, sketches, drawings, photographs, and price estimates are included. Data are included on pitching coal seams in the United States which potentially can be mined by hydraulic methods. Preliminary conclusions are presented, subject to revision as a result of economic feasibility studies to be completed during Phase III. An outline is presented of the final report, which will include economic studies and recommendations for research and development.

Jan 1, 1975

By E. G. Davis

A research program to evaluate the use of ceramic die inserts for ferrous diecasting was initiated by the Bureau of Mines at its Tuscaloosa Research Center. Data were compiled from the literature to characterize selected ceramic materials with respect to thermal properties, thermal-shock resistance factors, chemical reactions with molten metals, and processes required to form die components. With this information and a computer model developed for predicting temperature and stress patterns induced in ceramic materials during simulated ferrous diecasting operations, 27 candidate materials were evaluated. After evaluation, only 15 were considered for testing in an apparatus constructed to immerse the ceramic specimens in molten cast iron. Of these, 13 were tested in a modified testing unit in which cast iron was poured onto the specimens. These tests resulted in the selection of three materials for further testing: boron carbonitride (50 pct B4C:5O pct BN), boron nitride-grade M (BN-M, 40 pct BN, 60 pct Si02), and pyrolytic boron nitride (PBN). Tests in a gravity-flow die unit showed that PBN was thermal-shock resistant and could be used in diecasting of ferrous metals.

Jan 1, 1981

By Peter Radziszewski

"Typical mining operating costs can be divided between extraction (30 % to 70 %), comminution (30 % to 50 %) and separation (5% to 20% ). Cornminition process costs, representing an important contribution to mining operating costs, can themselves be divided roughly between energy (50 %) and linerkharge wear (50 %). Developing the ability of predicting mill wear will allow a more accurate estimation of wear costs. However, total media wear in a given ball mill grinding process is a product of three recognized wear mechanisms those being abrasion, corrosion and impact. Further, the contribution to total media wear of each of these wear mechanisms has not been well established. In this paper, a overview of abrasive, corrosive and impact wear mechanisms is made from which an empirical wear prediction procedure is proposed. Further, future directions in wear prediction research is outlined. INTRODUCTIONTypical mining operating costs can be divided between extraction (30 % to 70 %), comminution (30 % to 50 %) and separation (5 % to 20 %). Comminition process costs, representing an important contribution to mining operating costs, can themselves be divided roughly between energy (50 %) and lined-charge wear (50 %). With respect to wear, world wide steel consumption as grinding media is estimated at over 600 000 tomes per annum (Malghan, 1982). As such, developing the ability of predicting mill wear will allow a more accurate estimation of wear costs.Charge media wear prediction is actually estimated using the Bond abrasion test. From Bond abrasion index tests, it is possible to determine a wear rate using the following function:"

Jan 1, 2000

By Per-Anders Persson

This chapter deals with the concepts ofshock waves aud detonation waves together, because a detonation wave is really a shock wave, supported by the explosive reaction that the shock wave ignites and propagates as it travels through the explosive material. In the context of rock blasting, furthermore, there is justification for treating shock waves and detonation waves together, because the mechanical effect of the detonation in an explosive in a drillhole in rock is propagated and its effects are spread throughout the surrounding rock by means of a shock wave The overall performance of an explosive in rock blasting or any other application is of course dependent on how much thermal energy is released in the chemical reaction that occurs when the material detonates, and on how much useful mrc!~amcal work the reaction products can do as they expand. A major portion of this chapter deals with the prediction and measurement of explosive performance, a subject which is now still in a process of rapid development.

Jan 1, 1994

By P Moser, R Ivanova, F Ouchterlony

This paper describes model-scale bench blasts on mortar blocks with distorted drill hole patterns, done in 2013–2014 at the Montanuniversitaet Leoben. The main aim of the project is to see if the fragmentation from and the crack development behind blasts with and without drill hole deviations differ significantly.The dimensions of the test blocks were 660 × 280 × 210 mm (L × W × H). They were mounted inside a yoke that allows the blast waves to escape (see Figure 1). Three rows in each block were shot row by row with a nominal pattern of B × S = 70 × 95 mm. The same nominal delay 73 µs (ie 1 ms per m burden) between holes was used throughout. Blocks with holes of stochastic collar position errors (variations in burden or in burden and spacing but constant row volume) and systematically shifted collar positions (blasts with staggered pattern) were shot. The results were compared to reference blocks without collar deviations.Apart from the sieving analysis, the remaining bench faces were measured with a stereo-photogrammetric method and horizontal roughness profiles for each row were constructed. Sawing slabs from block remains after the third row and use of dye penetrant allowed the study of internal blast damage.The analysis showed that none of the changes in borehole collaring had a significant effect on the sieving curve (median). The sieving results confirm earlier findings though that the fragmentation gets finer with the number of rows shot, implying that blast damage from earlier rows has an influence on the blasting results. The tests demonstrated that the nequiv (equivalent n-value) has increased ten per cent by shooting staggered pattern and decreased by seven per cent by pattern with burden and spacing errors. For the rest of the patterns, the nequiv was roughly the same.The surface roughness investigation showed that a smooth surface was achieved by the staggered pattern. The other deviations patterns did not show a clear trend.The internal blast damage, which was measured in the remaining part of blocks, showed that out of eleven, four crack families seemed to be most affected by blasting. The comparison of the reference and distorted patterns showed that less cracks were apparent for the staggered pattern, ie less damage was introduced.CITATION:Ivanova, R, Ouchterlony, F and Moser, P, 2015. Influence of distorted blasthole patterns on fragmentation as well as roughness of, and blast damage behind, remaining bench face in model-scale blasting, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 693–706 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By L. A. Puchkov

The creation of automatic methane concentration control systems presupposes the use, as a control factor, of the air rates in workings and in mine as a whole. These systems present a complicated set of computers and software, including as major components the following - systems for detecting methane concentration and air rate; - telemechanics systems; - programs and means of data processing; - programs and means of air rate control in workings and main fan capacity control. The paper analyses the experience of development and exploitation of automatic methane control system in gassy mines of the USSR Donetz coal basin. Much attention is given to the study of technological factors influence upon aerogasdynamic process using automatic methane control system.

Jan 1, 1989

By L. S. Darken

It has long been recognized that the driving force in an isothermal diffusion process may be regarded as the negative gradient of the chemical potential (partial molal free energy) of the diffusing substance. This equivalence is exactly analogous to the fact that a mechanical force is the negative gradient of the potential energy. Nernstl in 1888 expressed these driving forces in terms of osmotic pressure, which is directly related to the chemical potential and free energy. The relationship between the diffusivity, mobility and activity coefficient of electrolytes in aqueous solution has been developed in great detail by Onsager and FuossZ with results which have recently been confirmed experimentally by Harned and Nuttall. This interpretation of the driving force in isothermal diffusion has been discussed earlier by the author,4,5 who pointed out4 that "for a system of more than two components it.is no longer necessarily true that a given element tends to diffuse toward a region of lower concentration even within a single phase region"; departure from the behavior of an ideal solution may be so great that the concentration gradient and the chemical potential gradient, or activity gradient, may be of different sign, thus giving rise to "up hill" diffusion. Since, in the discussion of that paper,4 some skepti- cism as to the validity of this conclusion was expressed, it seemed desirable to adduce experimental evidence to support it. Evidence of such diffusion in nonmetallic systems has already been reported. For example, it was observed by Hartley" in a series of experiments in which sodium chloride diffused from a saturated solution in about 5 pct acetone in water into a solution virtually free of salt but with essentially the same acetone concentration. Specifically, a wine glass containing crystalline salt and saturated solution was totally immersed in a large glass cylinder containing the acetone-water solution without salt; the difference in specific gravity was such as to prevent convection. Diffusion of salt was accompanied by a marked depletion of acetone in the wine glass; after 3 days the acetone concentration decreased from 5.37 pct at the top of the glass, where the salt concentration was low (0.93 pct), to 1.84 pct at the bottom where the salt concentration was high (21.64 pct). This "uphill" diffusion was found to be such that the partial pressure (activity) of acetone was essentially equalized throughout the system when a steady state gradient of sodium chloride was maintained. In order to demonstrate the existence of such diffusion in metals, a series of four weld-diffusion experiments was made. In these measurements pairs of steel of virtually the same carbon content, but differing markedly in alloy content, were welded at the end and held at 10500 C for about two weeks. Subsequent analysis showed that carbon had diffused so as to produce an inequality

Jan 1, 1949

By Marco Fellin, Van E. Komurka

Significant cost and construction benefits were obtained from performing a bi-directional static load test prior to construction on the Russell Street Bridge Replacement project in Missoula, Montana. The project utilized 24 drilled shafts to support two three-span bridges across the Clark Fork River. A bi-directional static load test (“BDSLT”) was performed on a large-diameter production shaft to demonstrate construction methods and confirm design assumptions. Subsurface conditions consisted of gravel with sand to a depth of 20 feet (6.1 m), underlain by gravel with sand, silt, cobbles, and occasional boulders. The test shaft had a nominal diameter of 5.9 feet (1.8 m), and an embedded length of 75.5 feet (23.0 m). The geotechnical engineer of record and the BDSLT test engineer collaborated to optimize the BDSLT jack assembly location within the test shaft, determine appropriate instrumentation locations, and plan the test protocol. The maximum applied test load was 10,707 kips (47.6 kN). Significant benefits were obtained from performing the BDSLT test, including 1) reducing each production shaft’s length by using a higher LRFD resistance factor allowed by having performed the testing, 2) using test-determined unit shaft resistances to evaluate and accept two production shafts that may have otherwise required remediation, and 3) providing the Montana Department of Transportation (and other designers) with increased presumptive unit resistance values in the gravel/cobble/boulder deposits. PROJECT DESCRIPTION The project was constructed in Missoula, Montana for the Montana Department of Transportation (“MDT”). The original bridge was approximately 60 years old, and consisted of a two-lane, 420-foot-long (128-m), four-span, riveted steel plate girder design spanning the Clark Fork River. The bridge was due for replacement owing to its deteriorated state, and the need for increased traffic volume capability afforded by a four-lane structure. HDR was the structural engineer. The preferred bridge alternate was a 459-foot- long (140-m), three-span, welded steel plate girder design. The new bridge was designed to meet current standards in accordance with MDT bridge design and AASHTO LRFD criteria. The new bridge was constructed in phases to keep traffic open during construction.

Jan 1, 2019

By H. A. Burgman

During in situ uranium mining using ammonium bicarbonate lixiviant, the ammonium exchanges with cations on the ore's clay. After mining is complete, the ammonium may desorb into post-leach ground water. For the particular ore studied, other elements (uranium and selenium) that are mobilized during the leaching process, have also been found in post-leach ground water. Laboratory column tests, used to simulate the leaching process, have shown that aluminum chloride can rapidly remove ammonium from the ore and, thus, greatly reduce the subsequent ammonium leakage level into ground water. The aluminum chloride has also been found to reduce the leakage levels of uranium and selenium. In addition, the aluminum chloride treatment produces a rapid increase in permeability.

Jan 1, 1985

By Song Guo

This paper will present the roof bolting finite element model developed at Jennmat Corporation used to study the control mechanism of the tensioned bolt system in the laminated roof with a large horizontal stress. In addition, it will discuss general support concepts pertaining to the tensioned bolt system which are derived 6om a series of studies in conjunction with the authors' working experience and extensive research efforts over the years.

Jan 1, 1997

By G. L. Oldright

In the previous study that has been made by the authors on smelting in the lead blast furnace,4 the furnace examined was one that operated under conditions that have been considered as more or less normal up to the present time. That is, although the furnace had in its charge a large portion of sinter, it also handled considerable quantities of naturally oxidized direct-smelting ore. Other ingredients, such as the limestone, were also added as fairly coarse material. The charges contained only about 24 per cent of lead. Until the changes incident to modern flotation began to take effect, many lead plants doing semicustom smelting in districts where competition was keen and much of the bonanza ore was exhausted were faced with the difficulty of obtaining enough lead to get a sufficient "lead fall" in the blast furnaces to clean the slag thoroughly and to separate it from the lead. A charge containing 24 per cent of lead would have been considered a very rich one by such smelters.

Jan 1, 1931

By G. Martino, F. Silva, Emmanuel Marchal

The definition of the chemical specification at multiple steps of the steelmaking process is very common in the industry. Among all elements and compounds, the Sulphur content has a special control over several stages, from the coal blend to the final products. Coke, iron and steelmaking specialists usually agree on a static specification for each intermediate product according to their local operation capacity and production cost. However, local decisions have major impacts on the whole production chain and many global factors should be considered in order to make the best decision. In this study, we will exemplify focusing on Sulphur content concepts and techniques that could be applied to many decisions in an integrated carbon steel plant. The variation of coal and coke prices, the desulphurization cost and time, the processes operating points, the hot metal rate at converters, prices and demands of low and ultra-low Sulphur steel grades are some of the integrated factors that are usually disregarded during the specification process. Mathematical modeling presents itself as a proper option into dealing with all the complexity that emerges from integrating all these processes' trade-offs and decisions. This study presents the results of a mathematical model that encompasses economic, thermal, chemical, and mass balances, physical quality of materials and productivity constraints of all processes of a hypothetical integrated carbon steel plant. It optimizes the iron and steelmaking process on a global and unique objective, dynamically defining targets for product quality to achieve the lowest steel production cost. This study details multiple market-plant scenarios in order to compare economic and operating results of a static vs dynamic specification practice.

Oct 1, 2019

Following unprecedented high market prices for tin and gold in 1980, significant advances in both mining technology and geological knowledge took place in northeast Queensland. These led to a rapid growth in mining activity. Production of these traditional metals reached levels comparable to those of the heyday of mining activity at the turn of the century.Although tin and tungsten prices have since collapsed, the geological knowledge gained and significant discoveries made during the boom, have created a potential for mine development when prices rise again. For other discoveries which are not workable because of treatment problems, future technological advances are required to raise their potential.

Jan 1, 1990

Tonnes per Hour - Kilowatts per Tonne Tonnes of Mesh Size Produced Kilowatts per Tonne of Mesh Size Produced - Terms often heard in discussions of grinding mill and grinding circuit performance. All of them only tell a partial story as none of them relate to a fixed reference or overall standard measurement. Tonnes per Hour gives the production rate relating it to a specified or desired rate set as a goal or objective. As such it does not take into account changes in feed size, product size or grindability. Kilowatts per Tonne which accurately stated is kilowatt hours per tonne brings power consumed into the picture which is related to grindability but still does not take into account changes in feed size and product size without which power changes do not necessarily reflect the complete effect of changes in Grindability.

Jan 1, 1978

By J. R. Pronovost, P. M. Tiley, D. W. Watts

"In 1980, J.S. Redpath Limited were retained by Dickenson Mines Limited to increase hoisting capacity from 700 to 1000 tons per day. The hoisting capacity of the present skip-over-cage system could not be increased significantly through design modifications. The other constraint was a minimum of disruption to the current production. The Dickenson shaft was a three-compartment shaft from surface to the 8th level and four compartments before the 8th to the 23rd level. The solution to the problem was to remove the services from the shaft man-way compartment from the 8th level to surf ace by drilling a borehole close to the shaft and rerouting those services through the borehole to the 8th level. A service compartment would then be available from surf ace to the 23rd level.By providing a service hoist in this compartment, the skip hoist would then be free to produce the 1000-tons/day required production. This paper will describe the search for the required hoist and the friction hoist selected for the project, and trace the design modifications required to modify the hoist to suit the Dickenson application and satisfy the Department of Labour safety requirements.The paper has been co-authored by J.S. Redpath Limited and C.L. Tiley Associates, who participated in the design changes and eventual commissioning of the hoist.IntroductionIn the earl y part of 1980, J.S. Redpath Limited conducted an engineering study on the Dickenson Mines Limited No. 1 Shaft materials handling system. The object of the engineering study was to investigate the possibility of increasing the tonnage hoisted from the No. 1 Shaft complex by adding a service cage compartment and service hoist. This would permit utilization of the existing hoisting plant solely for production hoi sting. A further consideration of the engineering study was that any alterations had to be accomplished with little, or no, interruption to production."

Jan 1, 1982

By Etienne Simode

When a fire breaks out in a mine access road, the hot smoke, due to heat transfer with the surrounding strata, progressively cools as it moves forward. Depending upon the resulting temperature distribution, the density of the fluid changes. This change will bring about: Firstly, a variation in the resistance of the roads, a static factor which will affect ventilation flow rates; Secondly, the appearance of extensive aerodynamic driving loads in the vertical and angled roads, a dynamic factor which can have an extensive impact on ventilation. Because of the appearance of combustion and distillation products, the vaporization of water, which wets the timbering and strata, and the expansion of gases, fluid rates can at times be increased tenfold in the area of the fire. These disturbances may cause explosions and ventilation inversions, which fill other working areas with smoke toxic to workers, and which can come up upon rescue teams from behind. Due to the progressive heating of the surrounding rocks, which will slow down heat transfer, these dangerous conditions change over time. At the Lorraine Basin Coal Mines, programs have been set up and employed for the computer simulation of these phenomena over time. They serve to anticipate or to understand such phenomena, the input data being the characteristic parameters of the mine ventilation network, the operating curve of the main fans, and measurements or estimates of some temperatures. Two programs, employed together, have been developed for these studies: One program, called RESO, calculates the distribution of flow rates and pressures throughout the entire network; The other, called FEUMIN, determines smoke temperatures over their entire path.

Jan 1, 1980

By H. L. Talbot, A. J. Hall, R. S. Young

A METHOD for the chemical differentiation of oxidized cobalt from sulphide cobalt in complex mine and concentrator products is of considerable interest to a number of mining companies throughout the world. At Rhokana Corporation, in Northern Rhodesia, for instance, cobalt is found chiefly in the sulphide form, as carrollite, but a cobalt oxide mineral also occurs irregularly underground, and relatively small quantities in the mill feed from time to time are difficult to detect it visually at the concentrator. The oxide cobalt is not amenable to the sulphide flotation process used at Rhokana and nearly all of it passes into the tailing. The occasional. period when high tailing losses occur with no appreciable change in the cobalt content of the mill feed has naturally been disturbing to the concentrator staff, who had no means of determining whether an influx of cobalt oxide ore or laxity on the part of the mill flotation operators were responsible for these results. A procedure for determining oxidized copper minerals, based on their solubility in dilute sulphuric acid saturated with SO2, has been used for many years in the copper industry. In the presence of cuprite the procedure fails, but otherwise it gives results quite satisfactory for control purposes. Empirical methods for determining oxidized forms of zinc and lead have been proposed, but they are generally unsatisfactory except in isolated cases of a very simple ore. No published procedure for determining oxidized forms of cobalt or similar elements has come to the authors' notice. EXPERIMENTAL PROCEDURE There are obviously two alternatives in attacking a problem of this nature. One is to employ a solvent for oxide cobalt that will not dissolve cobalt sulphide. The other is, to find a substance capable of oxidizing the cobalt sulphide to sulphate, which of course is water-soluble and can then be leached from the water-insoluble oxide cobalt. A representative sample of Rhokana cobalt oxide ore was secured, ground to minus 200-mesh, and carefully rolled to furnish a standard sample for this work. The analysis of this material gave the following percentages: Co, 0.55; Cu, 1.64; Fe, 2.98; S, 0.08; SiO2, 41.42; A1203, 11.20; CaO, 13.82; MgO, 3.80. As a source of cobalt sulphide a similar large sample of carrollite was prepared, analyzing Co, 23.36 per cent; Cu, 27.32; and S, 31.13. This sample of course was not pure carrollite, but was the best obtainable in quantities sufficient for a standard intended for considerable work. It was a pure sulphide sample, microscopic examination showing only carrollite, a little bornite, and gangue, with no visible sign of oxidized ore present. A wide variety of reagents were used in an effort to selectively dissolve oxide cobalt or to oxidize sulphide cobalt to

Jan 1, 1946

By G. M. Kintz

The data in this circular were taken from publications of the United States Bureau of Mines and published annual reports of the Wyoming State Inspector of Coal Mines. The tables included are complete only' with regard to major disasters; the Bureau of Mines considers a major disaster one in which 5 or more men are killed. However, minor disasters and gas ignitions recorded in these publications are included. It is known that a number of gas ignitions have occurred of which no record is available. This is one of a series of papers by the Bureau of Mines giving by States available data regarding coal-mine explosions. This information is published in condensed form in the belief that it may be of vale to mine operators in preventing future coal-mine explosions.

Jan 1, 1934

By R. A. Oriani

The partial molal volume of hydrogen is one of the parameters that describe the elastic interaction between the solute and the stress fields about inclusions, dislocations, and cracks. As such the partial molal volume is probably of importance in the elucidation of phenomena such as hydrogen embrittlement and hydrogen yield point. A knowledge of this quantity would also be helpful in thinking about the state of dissolved hydrogen in iron. However, because of the very low lattice solubility of hydrogen in iron the usual ways of determining the lattice expansion are not practicable. It is therefore of interest to apply the thermodynamics of stressed bodies to two sets of measurements of the effect of elastic stress upon the permeability of hydrogen in order to deduce a value of the partial molal volume, VH, of hydrogen dissolved in a iron. Beck ..' and previously de Kazinczy, observed that a uniaxial tensile stress increases the permeability of hydrogen in iron and in various steels. Beck et 01. employed Armco iron and A.I.S.I. 4340 steel, whereas de Kazinczy used a steel the composition of which was 0.13 C, 0.23 Si, 0.46 Mn, 0.006 P, and 0.038 S. Upon releasing the stress the permeability increment disappeared if the stress was below the elastic limit. Both investigators employed cathodic charging to introduce the hydrogen. de Kazinczy measured the permeation through a thin-walled tube by collecting the gas, whereas Beck et al. measured the permeation through sheets of various thicknesses by a very sensitive electrochemical technique. Both investigators measured the steady-state permeation at constant rate of hydrogen ion discharge, and Beck measured in addition the hydrogen diffusivity by a time-lag technique which is independent of the boundary conditions. Beck et al. and de Kazinczy found a linear relationship between log (J,/Jo) and the stress, where Ju/Jo is the ratio of the flux of hydrogen when the metal is under uniaxial tensile stress, a, to the flux under zero stress, for the same temperature and charging current. Beck et al. found in addition that the diffusivity of hydrogen is not changed by stress. Both investigators concluded that the observed change in permeability is due to an increase in hydrogen concentration, and furthermore that the increase in concentration is due directly to the thermodynamic effect of stress upon concentration. Accepting this assessment of the situation for reasons given below, one may use the equation3j4 in order to evaluate I/H, the partial molal volume of hydrogen. This equation is valid for the domain of a/E «¦ 1 (where E is the Young's modulus) and under the assumption that hydrogen expands the lattice isotrop-ically. From the data of Beck el al, one calculates V7H - 2.0 cu cm per g-atom, and from those of de Kazinczy one obtains 1.8 cu cm per g-atom. That the increase of concentration with stress is indeed of thermodynamic origin is attested to by the facts that the experimental results conform to the thermodynamic relation, Eq. [I], and that the results are the same whether a pure iron1 or each of two different steels172 is used. Neither of these facts wou1.d necessarily be expected if the effect of stress were, rather, to increase the ratio kn/k, of the kinetic factors of the following competing reactions: H(ads) — H (dissolved) Such a change of kinetics at the surface could be an alternative explanation of the effect of stress on the permeability. Although this writer does not deem this explanation to be the correct one for the reasons given above, it must be admitted that unambiguous proof that the phenomenon has a thermodynamic origin does not yet exist. Two kinds of experiments may be suggested. One is to plate a variety of metals on the input surface of the steel and repeat the stress experiment at a variety of hydrogen charging currents. The other is to employ somewhat thicker specimens in order to be able to apply uniaxial compressive stress. Eq. [I] shows that ln(c,/c,) depends on the sign of the stress, but it is difficult to see a physical basis for which k2/kl would depend on the sign of a. The present value of V^ in a iron agrees with phragmen's5 estimate, which he based on comparisons with the lattice expansion by hydrogen of titanium, zir-

Jan 1, 1967

By Chris Searing

An underground room-and-pillar limestone mine in Western Missouri acquired a new primary crusher, physically larger in every dimension than the unit it replaced. Due to the space constraints within the mine, resetting the truck dump and feeder to account for the larger crusher was unfeasible. Mine management determined the most practical solution was to increase the available area for the new crusher via blasting. Fitting the new crusher into the existing area required the removal of 6 feet (1.83m) of limestone from the original floor grade. Blast parameters had to be designed to effectively break this shallow depth of rock while controlling movement to avoid contacting the mine back 23 feet (7.01m) overhead, an apron feeder 13 feet (3.96m) above the southern extent of the area, and the original crusher which was still in place 15 feet (4.57m) from the first pattern. Additionally, work needed to be completed within an aggressive schedule set forth by the mine, allowing little room for delay or error. This paper will examine the efforts taken to safely and effectively blast the rock while protecting nearby equipment and integrity of the mine back. Subjects covered will include blast design, adjustments made to account for drilling and equipment-related constraints, and how seemingly small changes in load parameters can significantly impact performance with small-diameter blast design.

Feb 6, 2023