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By Susan M. Moore

The objective of this study was to evaluate the circumstances leading to fall from equipment injuries in the mining industry. Method: The 2006 and 2007 Mine Safety and Health Administration annual injury databases were utilized for this study whereby the injury narrative, nature of injury, body part injured, mine type, age at injury, and days lost were evaluated for each injury. Results: The majority of injuries occurred at surface mining facilities (~60%) with fractures and sprains/strains being the most common injuries occurring to the major joints of the body. Nearly 50% of injuries occurred during ingress/egress, predominately during egress, and approximately 25% of injuries occurred during maintenance tasks. The majority of injuries occurred in relation to large trucks, wheel loaders, dozers, and conveyors/belts. The severity of injury was independent of age and the median days lost was seven days; however, there was a large range in severity. Impact on industry: From the data obtained in this study, several different research areas have been identified for future work, which include balance and stability control when descending ladders and equipment design for maintenance tasks.

Jan 1, 2009

By J. Kronsteiner, E. Kabliman

"In the present work, we developed a Smart Process Control Toolkit (SProC) as a single simulation environment for the modelling of materials microstructure and its interaction with the local microchemistry during manufacturing of semi-finished products. This toolkit supports the data communication between sub-sequent processing steps such as casting, homogenization, pre-heating to deformation, etc. The thermo-kinetical simulation software MatCalc of MatCalc Engineering© GmbH was used for modelling of precipitation kinetics, such as formation of primary and secondary phases and their evolution. An import/export functionality of the precipitation distribution after each processing step was implemented to follow the parameter history. For the evolution of the local microstructure, a Mean Dislocation Density based Model (MD2M) was used. This model consists of two parts: (1) Flow Stress Model for calculation of the total strength during the deformation and (2) Static Recrystallization Model for account of the recrystallization and grain growth after the deformation. Validation was done by a comparison of experimental and simulated hot compression tests of a conventional AA2024 alloy.INTRODUCTION Optimization of a semi-finished products quality, elimination of typical production issues and cost savings are of a great demand for supply industries. However, it is currently not feasible to investigate the whole manufacturing process in detail at an industrial level. Concerning the processing parameters, one must understand the correlation mechanisms between macroscopic properties and microscopic material behavior. Numerical simulations, which could reproduce the real processing conditions on the one hand, and employ a physically based modelling on the other, are thus attracting a lot of attention and their power is no longer under the question. It is well known that e.g. the total strength of the material depends directly on distribution of precipitation particles and mean grain size. In order to follow their evolution during the manufacturing process beginning from casting to the final semi-finished product, and thus be able to control the process, one should use the techniques which would at the same time reproduce experimental processing conditions of real size objects (e.g. sheet, extrusions, etc.) and account for the evolution of a local microstructure and its interaction with the microchemistry. A commonly used approach is to implement a model which would describe the evolution of microstructure based on a dislocation density approach into a user defined material subroutine of some finite element solver. There is, however, also an influence of the manufacturing process by e.g. the evolution of a local microchemistry (formation, transformation and dissolution of the precipitation particles) on the microstructure and thus the macroscopic material behavior. Thus, a method is needed to follow the variable history through the simulation of the whole production chain."

Jan 1, 2018

By Z. H. Xu

Flotation as a versatile multiphase separation technology has been practiced for more than 100 years in coal and mineral processing. Although effective over a wide particle size range, flotation for mineral separation has been largely understood as an art rather than a science, treating the flotation system as a black box. With rapid progresses in microscopic technologies coupled with high speed video recording and imaging analysis, ingénue design of visualization systems has made it possible to understand fundamentals of flotation by visualization. In this presentation, the state of the art where visualization plays a key role in advancing our fundamental knowledge of flotation will be reviewed. The areas to be covered include characterization of bubbles (shape, size, velocity and sol-ids loading) in a running flotation system, either column or mechanical cell, bubble-solid attachment, hydrodynamic cavitation in fine particle aggregation and flotation, froth imaging, particle-particle interactions by zeta potential distribution measurements, and heterogeneous gas nucleation by atomic force microscope imaging. With these capabilities, critical roles, good or bad, of ore proper-ties, solution chemistry, slurry hydrodynamics, and operation variables in flotation systems as applied to mineral, coal and oil sands processing were clearly elucidated.

Jan 1, 2014

By J. Borlée, J. P. Birat

"ULCOS is a large European program launched by a consortium of 48 partners led by arcelor and a core group of steel producers, which was started in 2004 with the support of the European Commission (RFCS and 6FP programs) to search for breakthrough process routes that could, in the future, when fully developed, demonstrate the potential of large cuts in CO2 emissions for the production of steel from iron ore.The timescale for commercial implementation is the post-Kyoto period, i.e. 15 to 20 years from today.The present paper presents the progress made in the first 2 years of this 5 year program.Several concepts are being investigated in parallel: re-engineering of the blast furnace to incorporate CCS in the core of the process, smelting reduction with CCS capabilities, massive use of natural with more CO2-lean technologies than today, hydrogen steelmaking and electrolysis of iron ore. Biomass also offers very attractive perspectives, while raising major challenges.None of these new steelmaking routes are ""no regret"", which means that the European Steel Industry has already incorporated extensive energy saving steps in its present process technology and is working as close to thermodynamics limits as is cinematically and economically feasible, and that decreasing GHG emissions further means introducing breakthrough technologies that add significant costs to the operation of the Steel Mills.They will also require much research and development before they can be proven and implemented commercially.IntroductionThe ULCOS program (Ultra Low CO2 Steelmaking) is a large research program of the European Steel Industry and its industrial and research partners launched in September 2004 to investigate steelmaking technologies that have the potential of reducing the CO2 emission of the Industry by a large amount, typically in the order of 50%. It is supported in the European Community by the 6th Framework and the RFCS programs through 3 complementary European projects [1,2].The present paper summarizes the work carried out in the phase 1 of ULCOS and in framing the program of phase 2."

Jan 1, 2008

By W. H. Oppelt

Two exploratory gasification tests of limited duration were made at Dorsten, Germany, with strongly coking Pittsburgh-bed coal in a fixed-bed pressure gasifier. To reduce the coking potential of the Arkwright coal, ash refuse from the gasification of Dorsten (Leopold) coal was added to pre-pare coal-ash mixtures which contained nominal 20 and 30 percent ash. Major adjustments in principal process variables at the beginning of the tests resulted in wide fluctuations in the rate of gas production, gas offtake temperature, and carbon dioxide concentration in the gas. Because a higher steam-to-oxygen ratio was used in the tests than was customarily employed for routine operation, gas production was lower, and steam and oxygen consumption was higher. After operating conditions had been improved, the results of a 12~-hour-long test with addition of sufficient ash refuse to obtain a nominal 30 percent ash concentration indicated that operation is feasible with this coal mixture using a fixed-bed pressure gasifier equipped with a rotating coal distributor and stirrer.

Jan 1, 1966

By Jeff Kravitz, John Kovac, Wayne Duerr

Since 198 1, Federal mining law has required every miner working in underground coal mines to have a self-contained, self-rescuer (SCSR) available for use in emergencies. Some miners that have escaped from fires using SCSR's have reported that they had to remove their mouthpiece to talk during escape, thus compromising the protection afforded by the SCSR. If miners are in thick smoke during an escape, they may have trouble if they become separated from their group. The visibility can be so poor that separation can occur at very short distances. To address this problem, two-way, FM radios were built into the SCSR to improve the ability to communicate should the miners be separated over relatively short distances. The SCSR prototypes built and used in testing included both two-way radios and improved mouthpieces. Comparison studies were conducted using combinations of the new and old mouthpieces, and trials using the radio or not using the radio. Subjective judgments indicate that communication ability was significantly improved using the prototype devices. Communication problems have been encountered with teams using standard voice- powered communications systems. Mine rescue teams perform rescue and recovery work under extremely hazardous conditions, life lines are a standard part of mine rescue team equipment. Leaky feeder and quasi-leaky feeder life line communications systems provide significant improvements over standard systems. These improved systems allow all members of the team to communicate with each other and their base of operations. These innovations will significantly enhance the safety of mine rescue operations.

By K Katov, S T. Ciutina, J D. Pyke

The development and optimisation of mill discharge trommels, and specifically their screen panels, can be a neglected area in the complex operation of a minerals processing plant. Trommels are often seen as a ‘set and forget’ component of the circuit, even when significant improvements in design are possible. A redesigned trommel and panel arrangement has simplified and added flexibility to the plant shutdown schedule at Kirkland Lake Gold’s Fosterville Gold Mine in Central Victoria.Prior to the trommel redesign, the shutdown frequency was determined by the expected life of trommel panels at 14 weeks. Planning downtime around this item led to higher maintenance costs and wastage in other areas of the plant. This was particularly disruptive with respect to mill lining change-outs, which were forced out of sequence and often changed prematurely. Fosterville Gold engaged Metso Australia Limited to propose a new trommel design that would better suit the site. The new trommel was engineered to balance simplicity with performance, and incorporated the latest in Modular trommel design, modelling and Finite Elemental Analysis (FEA). Modular Panels and Spiral Segments achieve significantly longer life, and allow in-situ panel change outs, flexibility in spiral and dam positioning and an improvement in classification performance. Trommel panel life has significantly increased from 14 to 71 weeks in the trommel’s highest wearing zone. The new design has eliminated some risks to safety by reducing manual labour and lifting requirements during shutdowns. There is greater flexibility with respect to shutdown requirements. The site is no longer at the mercy of trommel panel lifespan, which allows for better optimisation of wear components in the circuit. CITATION:Pyke, J D, Katov, K and Ciutina, S T, 2018. Modular SAG mill discharge trommel for maintenance flexibility and performance, in Proceedings 14th AusIMM Mill Operators' Conference 2018, pp 613–620 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

By Ho-Sang Sohn

Copper concentrate particles of 200 to 300 mesh size were fed from the top of vertical reaction tube of 2.8 cm ID and 2 m long with an O2-N2 gas mixture. The reaction tube was heated to 900 K to 1100 K. The residence time of the particles in the reaction tube was about 1 second. The copper concentrate particles were very rapidly oxidized and melted down during their descent in the reaction tube. The particle temperature were calculated by combining an un-reacted core model, mass transfer between gas and particles, and heat transfer between gas, particles and tube wall. The particle temperature reached its maximum within a distance of 60 cm from the top of reaction tube, and it attained at 2000 K at higher oxygen partial pressure. The most particles were melted at the oxygen partial pressure above 20 kPa. An appreciable amount of As, Sb and Pb in the concentrate particles was eliminated in the upper portion of the reaction tube and the rate of elimination are explained by the rate of mass transfer of volatilized species through the gas film on the particle surface and the variation of particle temperature during the descent.

Jan 1, 2006

By Joseph J. Garbarino

World demand for phosphate fertilizers has grown at an average annual rate of between 6 and 7 per cent over the past 10 years. In the next five years world demand is expected to at least continue to grow at that rate. Also, the continued need for improved fertilizer materials, coupled with improvements in fertilizer technology, have placed added importance on the use of phosphoric acid as a starting material for high analysis phosphate fertilizers. This demand for phosphates and phosphoric acid must be met by increased production. Increased phosphoric acid production also brings increased demand for chemical processing agents including foam control agents. We at Arizona Chemical Company have closely followed the phosphate fertilizer industries needs. Arizona Chemical Company has been an important supplier of flotation agents to the phosphate industry, primarily from its Panama City, Florida plant. We have also recognized the important role played by foam control agents in the wet-phosphoric acid manufacturing process. We feel it is the responsibility of the supplier of foam control agents to offer specifically designed products that are highly efficient and economically useful. The difficulties involved in evaluating foam control agents in the plant are recognized. Owing to the large number of variables present, and the difficulties in controlling these variables during the period of a plant trial, plant evaluation of a large number of foam control agents is impractical. We have attempted to develop a practical, reproducible, and yet rapid, laboratory evaluation test which correlates with plant performance.

Jan 1, 1966

By C. Chamberlain

The Telfer Project, Australia's newest, largest, and most remote gold mine treats 480 kt of ore grading 9.33 g/t derived from open pit mining operations. Coarse gold is recovered by gravity concentration with jigs and further upgraded wing a continuous strake launder table, and by amalgamation. Remaining gold is recovered in a cyanide circuit and incorporates flat bottom air agitators, counter current decantation (CCD) and stellar filters for clarification and precipitate removal. Overall recovery into a bullion averages 98%. The mill incorporates a number of design features that allow for maximum efficiency and security wing a minimum number of operation personnel.

Jan 1, 1983

By D. P. Desmond

The U.S. Bureau of Mines investigated leaching automobile catalysts with sodium cyanide (NaCN) solutions at high temperatures to recover platinum-group metals (PGM). The feed was virgin monolith rejects, used monolith, and used pellet catalysts. Leaching with a 5-pct NaCN solution for 1 h at 160° C dissolved over 97 pct of the PGM in the virgin monolith, over 85 pct of the PGM in the used monolith, and over 90 pct of the PGM in the used pellet catalyst. Over 99.8 pct of the dissolved PGM was recovered as a precipitate by heating the solution to 250° C for 1 h in an autoclave. The cyanide complexes were decomposed, and free cyanide was destroyed.

Jan 1, 2010

By David La Rosa, Matthew Craft, Lee Julian

Rules of thumb are ubiquitous in blasting. Typically presented as ratios, they assist with the initial assumptions of blast design and provide guidance for best practice. Examples include the blast hole diameter as a function of bench height, and stemming length as a function of hole diameter. The most common rules of thumb were developed in mid-20th century and it is important for a blaster to know if the rules of thumb are still reliable.

Feb 6, 2023

By Christopher B. Vierrether

Reflected light microscopy, scanning electron microscopy, and electron microprobe analysis have been used to determine the mineralogy, phase percentages, grain sizes, textures, and partitioning of cobalt and nickel in two cobalt-nickel rich mattes. These mattes were prepared by submerged bath smelting from a flotation cobalt-nickel concentrate containing 4.8% cobalt and 5.1% nickel. The concentrate was derived from lead-zinc-copper ores of the Viburnum Trend, southeast Missouri. Eleven discrete phases have been distinguished in the matte samples, three of which contain significant cobalt and nickel. The phases are cobaltpentlandite ((Co,Ni,Fe)9S8), cobalt-bearing marmatite ((Co,Fe,Zn)S), shandite (Ni3Pb2S2), bornite (Cu5FeS4), galena (PbS), magnetite (Fe3O4), pyrrhotite ((Co,Ni,Cu,Fe)1_xS), chalcocite (Cu2S), CuPbS phase, metallic copper, and metallic silver. Cobaltpentlandite and cobalt-bearing marmatite- occur in coarse grains (50 µm) that would separate readily during grinding, but shandite contains small (5 - 10um), abundant inclusions of pyrrhotite and bornite.. Some mineralogical variations between the two matte samples were due to differing degrees of oxidation during smelting,

Jan 1, 1988

By L. Rahnenfuhrer

"Hoisting equipment is a vital part of the operation of most producing mines. Safety and continuation of operation necessitates ultrasonic preventive maintenance inspection on a regular basis.Ultrasonic inspection of hoist shafting, brake linkage and draw bars will contribute to a safe operation and can prevent costly unscheduled shut-down time. With little dismantling, ultrasonic testing techniques can be applied to. determine the conditions of shafting and linkage in hoisting equipment, mining equipment and other miscellaneous mining equipment."

Jan 1, 1969

By T. D. Yensen, R. K. McGeary

While we have not been primarily interested in the determination of hydrogen in the alloys that we have been dealing with, we arc very glad to cooperate in this symposium on sampling and analysis for hydrogen, as we can never tell when we may find it desirable to perform such analyses. Vacuum-fusion Method Our primary interest has been oxygen and carbon, in the analysis of which we think we have become very proficient. Our oxygen-analyzing apparatus is of the vacuum-fusion type (Fig. I). This employs Dr. Derge's type of graphite cruciblc and slit sleeve, with graphite powder between, for heat insulation (Fig. 2) supported in a clear quartz bulb connected to the Pyrex glass system by means of a carefully made ground joint. The latter is kept cool by means of an air blast from a circular copper tube surrounding the joint. The crucible is heated by high-frequency induction and the 24 one-gram samples successively dropped in from the horizontal tube at the top by means of a magnet. The gases (H², CO, N2) are passed through the CuO catalyst held at 300°C. (converting CO to CO²I, H² to H²O) and the CO² and H2O frozen out in the main liquid air trap, while N2 is evacuated. At the end of the operation, shown by the termocouple vacuum gauge indicating a pressure of 10-4 mm. Hg, the analyzing system is isolated by means of the mercury cutoffs, and the CO² and H2oare determined by transferring first the CO2 and then the HzO to the small trap at the right. This is done by placing liquid air on the small trap, then removing the liquid air from the large trap for one minute, which is sufficient to evaporate the Co2 and transfer it to the small trap. The CO2 is then measured by closing the mercury cutoff at the extreme right, removing the liquid air from the small trap, warming it up to room temperature, adjusting the mercury level to one of the three marks on the glass tube (corresponding to three volumes calibrated so that I mm. Hg of CO² corresponds to 0.001, 0.002 and 0.010 mg., respectively, of oxygen) and finally reading the pressure. Liquid air is now replaced on the large trap, the mercury cutoff at the right opened, transferring the COz back to the large trap. The cutoff to the pumps is then opened, the liquid air removed for one minute, thus evacuating the CO², the pumps again cut off and the H2 O transferred to the small trap as before, except that it is now necessary to warm the large trap with a Bunsen burner in order to evaporate all the H2O. This requires about three minutes, as the inner tube must be heated to at least 100°C. In order to prevent H²O from being adsorbed on the glass walls, the glass tubing beyond the catalyst (i.e., to the right in Figs. I and 3) is kept heated to about 150°C. by means of Nichrome ribbon as shown in Fig. 3. For a I-gram sample the above amounts of oxygen (0.001, 0.002 and 0.010 mg.) 35

Jan 1, 1945

By Walter Renton, Ingalls

SEVERAL years ago I became interested in computing the historic lead production of the United States, and the mines, or mining districts whence derived. This led me subsequently to an examination of the recent records- of some of the American zinc-mining districts and of the great lead and zinc-mining districts in foreign countries. My present article thus will be to a large extent an addendum to my book, "World Survey of the Zinc Industry," which was published by the Mining and Metallurgical Society of America in 1931, and to which I shall refer hereinafter as Ingalls, op. cit. Throughout this article all expressions of tons mean the ton of 2000 lb. Where original data have been given in metric tons I have converted in many instances simply by multiplying by 1.1, and long tons by 1.12. I have not been fussy about digits and have rounded off many figures. These may explain differences between data herein and other data.

Jan 1, 1946

By Charles H. Thurman

THE data given herein were first included in a paper read before San Francisco Section, A.I.M.E., in October 1940, and are applicable to conditions existing until the gold-dredging industry was temporarily suspended by Order L-208 of the War Production Board in the fall of 1942. There is no doubt that gold dredging will be resumed at some future time, and on a scale equivalent to that of the recent past, but the cost data given herein will be affected to some extent by conditions existing at that time, and by labor rates and prices for materials then prevailing. DRAGLINE VS. BUCKET-LINE Dragline dredges should not be considered for mining large areas of low-grade placer gravel. In such deposits a dragline operator with a 5-cu. yd. dragline would be doing well to maintain an operating cost of 6¢ per cu. yd., while a bucket-line dredge with buckets no smaller than 7 cu. ft. each could be operated for less than 4¢ per cu. yd. By operating costs, I mean only field charges such as direct operating and overhead, which includes pay roll, power or fuel, repairs and maintenance, insurance, land clearing and general expenses, but not including such items as prospecting, depletion, depreciation of equipment, State Franchise Tax, or Federal Income Tax. It is not to be inferred that a dragline dredge cannot make a profit in small low-grade deposits, where the extent of a given property is too small to justify the use of the more costly bucket-line dredge. Consider small deposits of 10¢ gravel. It is obvious that a deposit of 5,000,000 cu. yd. of such gravel could not be mined profitably with a new, 7-cu. ft. bucket-line dredge, costing upward of $250,000, and with operation costs at 4¢ per cu. yd., paying for land or royalties out of earnings; nor could a new combination 5-cu. yd. dragline and washing plant, worth upward of $125,000 and working at 6¢ per cu. yd., make a profit. These comparisons are made because the hourly capacity of both types of dredges is nearly the same. The bucket-line dredge averages 93 per cent operating time in 24 hr., as against about 80 per cent or less for the dragline dredge. These percentages of operating time are important factors in making a comparison of this kind. The bucket-line dredge has greater efficiency in cleaning soft bedrock, usually characteristic of dredging deposits, and a 7-cu. ft. bucket-line dredge can be efficient to a total digging depth of 80 ft. (65 ft. below water, 15 ft. above). A 5-cu. yd. dragline dredge is limited to about 30 ft. of efficient digging depth, 5 ft. of which could be above water. In small areas of gravel not over 30 ft. deep, a 5-cu. yd. dragline dredge could make a profit where a bucket-line dredge could not do so. As an example: compare the dredging of 3,000,000 cu. yd. of 10¢ gravel 28 ft. deep with bucket line, costing 4¢ per cu. yd., and with a 5-cu. yd. dragline dredge operating at 6¢ per cu. yd. Because of small yardage, new equipment is not warranted, therefore we can assume that the dredges of both types are used equipment, fully depreciated. Dismantling, mov-

Jan 1, 1945

By J. A. Mielczarski

An infrared external reflection spectroscopy (ERS) technique has been developed to study the structure of the adsorbed layer of ethyl xanthate formed on copper and natural chalcocite (Cu2S) electrodes. The infrared technique provides information concerning the identity and the orientation of the adsorbed species at monolayer and submonolayer coverages. X-ray photoelectron spectroscopy (XPS) studies have also been conducted to monitor changes at the surface. The results show that the first oxidation peak on the voltammograms for copper and the first and second oxidation peaks on the voltammograms for chalcocite electrodes are not characteristic of a chemisorbed xanthate species, which is contrary to what has been reported by other investigators previously. As evidenced by the XPS spectra, there was an increase in sulfur relative to copper on the copper and chalcocite surfaces over the potential region where these oxidation processes occur. The spectroscopic results (ERS, XPS) indicate that the second and third (on copper) and the third and fourth (on chalcocite) oxidation processes recorded on the voltammograms are due to cuprous ethyl xanthate adsorption. The orientation of the adsorbed molecules of xanthate was observed to change with increasing surface coverage. Therefore, it has been concluded that xanthate can modify the copper and chalcocite surface in two different ways. At lower potentials, xanthate is a source of sulfur for the formation of a sulfur-rich surface. At higher potentials, an adsorption layer of cuprous ethyl xanthate is formed. Contact angle measurements show that both of these processes of xanthate surface modification lead to an increase in the hydrophobicity of the substrates.

Jan 1, 1990

By Donald L. Everhart

The National Uranium Resource Evaluation (NURE) program since 1974 has been designed to accomplish a comprehensive coverage of the 48 conterminous States and Alaska to evaluate the total uranium resources of the USA. This resource base includes both ore reserves and potential resources. Techniques include airborne radiometric and magnetic surveys, by hydrogeochemical surveys, surface and subsurface geologic studies, and the application of improved geochemical exploration and resource assessment methods (Table 1). In brief review, the current status of each of the five basic NURE elements is as follows: In the Aerial Surveys, in the conterminous 48 states, the flying has been completed in 89 of the NTMS 10 x 2° NURE quadrangles, and is in progress in 39 more (Figure 1). Complete quadrangle data have already been open filed for 46 additional quadrangles, for a total of 174 for which data gathering is either complete or in progress. In Alaska, the flying is complete in 42 quadrangles, and quadrangle data have been open filed for 34 more, for a total of 76 (Figure 2).

Jan 1, 1979

By A. J. Farstad

Electromagnetic location tests were conducted by Westinghouse Georesearch Laboratory under the sponsorship of the U. S. Bureau of Mines at the Galena Mine in Northwest Idaho. The objectives of these tests were to evaluate the performance in deep hardrock mines of an electromagnetic location system previously developed for the U. S. Bureau of Mines for use in coal mines. Tests were performed using long wire transmitting antennas as well as horizontal and vertical transmitting loops at depths from 3000 feet to 4300 feet (995 metres to 1311 metres). Surface measurements were made of magnetic field strengths, field strength patterns, and background noise, These measurements were compared with theoretical predictions for a homogeneous ground conductivity consistent with this particular geologic region. The results of the Galena Mine tests show conclusively that light-weight portable electromagnetic transmitters can be effectively utilized to provide uplink communications and location information in deep hardrock mines. Typical metal mines such as Galena utilize deep vertical shafts for mine access and haulage and these shafts produce signal enhancement at the surface from underground electromagnetic transmitters by as much as 30 dB. The results also show that horizontal wire configuration can serve as an effective alternative to the magnetic dipole antenna as a rescue transmitting antennaa and can be used when the required area is not available for deployment of a loop antenna. In spite of the electromagnetic coupling effects observed in the vicinity of the shafts, the location accuracy obtained still provided useful information on the general location of the transmitting sources involved. The horizontal wire location accuracy was within 46 metres, while the vertical magnetic dipole and the horizontal magnetic dipole accuracies were within 152 metres and 110 metres respectively at a 915 metre depth.

Jan 1, 1973