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By Kanaan Hanna, Mir Heydari, Dipack Sengupta, Gordon B. French

This article, first in two, discusses the safety, technology, and economic impact of eliminating safety fuse from metal and nonmetal mines. It is based on operation and manufacturer experience over the past five to 10 years. If safety fuse is eliminated, blasting accidents will decrease. There should be only limited technical impact. The economic impact will be minor for large operations, but will be significant enough for small operations that some will likely close rather than change. Most mines would probably change over, with the market absorbing the extra cost. Elimination of safety fuse would obviously have an impact on the two safety fuse manufacturers. One manufacturer produces an alternative system and would be unaffected. The other, however, will be affected if it does not change its production strategy. From the standpoint of employment and production, abolishing safety fuse will have a very minor effect on the mining industry as a whole. This article discusses the current use of initiation systems, incidence of accidents, and fuse consumption. Next month, part two describes the economic impact of eliminating safety fuse on the mining industry and manufacturers, and presents an economic analysis of the initiation systems. Introduction The objective of this study was to determine the safety, technology, and economics of eliminating safety fuse from metal and non-metal mines. There were five major tasks: • Studying initiation-related accidents from the Health and Safety Accident Center (HSAC) data file; • Visiting mines, collecting relevant data, and interviewing mine operators; • Establishing the trend in safety fuse consumption and type of users; • Corresponding with fuse manufacturers, domestic and foreign; and • Developing an idealized blasting cycle to evaluate the possible economic impact of safety fuse elimination on large and small operations. More than 450 explosives-related accident reports have been screened and analyzed by investigating HSAC files. These reports have been studied and classified according to the four initiation systems, size and type of mines, and time and frequency of accidents from 1973 through 1978. Twenty-four mining operations were visited to obtain supporting data for this study. Mine personnel were interviewed to determine attitudes about safety fuse and alternative detonating systems. This article decribes available data. It also attempts to provide information that can be used to answer such questions as: • Is the use of safety fuse (an antiquated detonation system) a major cause of mining accidents? • Are alternative systems available that could result in fewer in-juries to miners?

Jan 1, 1984

By R. L. Loofbourow

In addition, dewatering is seldom complete because of aquifer irregularities and an operator's natural reluctance to start a costly study program, provision of wells, and pumping long before mining begins. In some cases, water pumped from the mine is used in treatment. This use is suggested as a way to minimize effluent effects. The true objective of water supply, however, is to provide water of satisfactory quality at the rate needed and at the least cost. Pumping to permit efficient mining isn't likely to coincide strictly with that objective during much of the mine life. The history of mining is replete with efforts, some highly ingenious, to control mine water. Methods that might be used to tackle the challenge Operators of wet mines are faced with both obvious pumping costs and some unquantified water costs. It is possible to reduce both types of costs, generally by reducing inflow to the mine area-and a surprising number of methods have been used to do this. Both the obvious and unquantified costs at a wet mine are now multiplied by: •Increasing costs of power, labor, equipment, and supplies; •Work at greater depth and, in some cases, in wetter ground; •Increasing use of hydraulic fill; and •Increasing zeal in protection of surface and groundwater quality. There are many easily overlooked ways in which mine water may add to hazards and costs (Loofbourow, 1973). As to their importance, consider these examples: •Increased maintenance of equipment, tires, roads, and ditches; •Increased difficulty in supplying proper ventilation; •Increased hazards from ice and use of electric power; • Loss of friable ore minerals and fines stuck on conveyances and in chutes; •Need to use more expensive methods and explosives; and •Decrease productivity of equipment and personnel. Expressed as cost per unit of ore produced, these unquantified costs can be high. Ideally, advanced dewatering should eliminate most of them. Advanced dewatering is a highly effective water control method, but it isn't problem free. Mines in ideal ground for advanced dewatering are most unusual. To be effective, water must be pumped at a comparatively high rate for a period longer than the life of the mine, from below mining depth. Effects on both groundwater and surface water are likely to be drastic. of increasing mine water costs are listed in the accompanying table. Ten groups of possible methods are cited. Most have been demonstrated in mining; two are known from industries other than mining.

Jan 12, 1980

By Toyohisa Fujita, Mitsuo Mamiya, W. Tai Yen, Hiroki Kogita

Green algae taken from an acidic mine drainage and blue-green algae taken from an alkaline hot spring stream were collected and tested for their ability to recover or remove dilute metal ions. Experimental results demonstrated that unwashed blue-green algae and washed green algae effectively adsorbed base metals ions and eluted them at pH 1. It was also found that washed and dried algae adsorbed precious metal ions more effectively than unwashed algae. For example, the washed and dried blue-green algae was capable of adsorbing 0.31 kg of gold per kg of algae. The gold from tetrachloroaurate solution which was adsorbed on washed blue- green algae was found to change to a metallic state following initial metal binding. In the case of a dilute gold complex solution leached with thiourea, only a small amount of gold could be captured by algae. Further experiments were conducted on components of the algae, such as alginic acid, agar, cellulose and chitin and mixtures of these components, in order to determine their contribution to metal adsorption characteristics. However, a mixture of these two components demonstrated both good adsorption and desorption characteristics indicating an interaction between the individual components.

Jan 1, 1995

By W. Kennedy Gauger, Tim Oolman, Debra L. Bisson

Radian Corporation (Radian) has developed a process for treating Cr(V1)-contaminated soils that involves: 1) soil washing using an alkaline extractant, 2) biological reduction of Cr(V1) to Cr(II1) in the soil wash effluent, and 3) separation of Cr(II1) from the biologically treated effluent. Biological treatment has been demonstrated for immobilized cells in fluidized bed bioreactors and suspended cells in slurry bioreactors for Cr(V1) levels ranging from 30 to 600 mg/L which have been reduced to non-detectable levels (< 0.05 mg/L; method detection limit [MDL]) during hydraulic residence times as short as 12 hours. The resulting Cr(II1)-containing effluent stream was chemically treated to precipitate chromium hydroxide for removal by physical separation.

Jan 1, 1995

By Dean S. Rogers, R. A. Metz

The conventional role or exploration, development and grade control is examined in this paper within the context of the many and varied types of gold occurrences which are found at the Dome Mine. The account endeavours to trace the development of the total ore evaluation process, from diamond drilling through the exploration and drifting stage, to the stope development and planning phase of the operation, with an emphasis on sampling techniques and assaying procedure. The concept of ore reserve calculations is developed and the practice of grade control is described. In summary, the role of the geological, the engineering and the mining department is presented in relation to the contribution each has to make in the evaluation and exploitation of ore from one of the world&apos;s classic Archean gold deposits.

Jan 1, 1992

By J. D. Megeath

In the October 1988 issue of MINING ENGINEERING, J.G. Wargo&apos;s paper discusses the problem of measuring and analyzing permeability, porosity, and fracturing in its relation to in situ leaching of disseminated gold deposits. This paper made me suspect that a good deal of information and analysis may have been done on this in reverse in the petroleum industry. Some years ago, I was involved with the US Geological Survey in a project studying the Devonian shale of the eastern United States. The Devonian shale contains gas deposits that have the characteristic of slowly seeping gas over a very long time period. To understand the rate of seepage, considerable study and analysis was done on the interrelationship of porosity, permeability, and fracturing. The purpose was to find out how and why the gas could escape in the varying rates found in wells throughout the formation. It strikes me that this is the mirror image of the problem faced in attempting to understand how a cyanide solution will travel into a deposit. To possibly invent our own personal wheel a little faster, I would suggest a look at the work done on the Devonian shale. Some of it is contained in "Estimates of Unconventional Natural-Gas Resources of the Devonian Shale," US Geological Survey Open-file Report 82-474, 1982, and "Additional Data for the Estimates of Unconventional Natural-Gas Resources of the Devonian Shale," Open-file Report 83-676, 1983. It is entirely possible that some of this previous work could shorten the time required to make in situ leaching of disseminated gold deposits a profitable technique.

Jan 1, 1990

By P. C. Rath, K. Srinivasa Rao, T. Subbaiah

A process for the recovery of nickel from ammoniacal leach liquor through an integrated solvent extraction-electrowinning (SX-EW) flowsheet was tested on a pilot-plant scale. The objectives of the test run were to assess the reproducibility of laboratory-scale data in a large-scale continuous operation and to study the build up of impurities in the system. The pilot-plant test was conducted under operating conditions that were established from the laboratory-scale experiments. Efficiencies of nickel extraction, nickel stripping and ammonia scrubbing were comparable to the laboratory-scale values, whereas the current efficiency during nickel electrolysis was observed to be less than the laboratory-scale values. The electrodeposited nickel contained some cobalt and iron impurities.

Jan 1, 1998

By J. M. Mutmansky

The objective of this study was to produce coal dusts for laboratory research that are similar to mine airborne dusts in chemical, mineral, and physical characteristics. In this investigation, dusts produced in laboratory comminution equipment were compared with continuous-miner-generated dust for similarity in physical characteristics. Mining-generated dust (MGD) and bulk coal samples were first collected from an underground mining operation. The bulk samples were then processed in the laboratory for size reduction in various comminution devices. These laboratory-generated dust (LGD) samples were subjected to size and microscopic analyses. Physical properties of the dust particles were studied on a scanning electron microscope (SEM). A numerical shape coordinate system (SCS) was also developed and applied to 300 particles from each sample. Statistical analysis was then performed on the collected data and characteristics of the MGD and LGD samples were compared.

Jan 1, 1995

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Sublevel long-hole stoping is a large tonnage stoping method that may be used when there is a fairly com¬petent straight vein and the walls will stand with little support. It is a method that lends itself to a high degree of mechanization. Two variations of the method are illustrated in Figs. 1 and 2. Fig. 1 illustrates development typical for a 3-m (10-ft) wide vein. Initial development is by a lower haulageway, which may be either track or trackless. A two-compartment raise is driven in the footwall of the vein. At about 12 m (40 ft) above the main head¬ing, a footwall scram is driven and pockets through the vein are cut on 7.6-m (25-ft) centers. Each of the sub¬level blocks selected in this example is 45.7 m (150 ft) long by 12 m (40 ft) high. Stoping starts at the slot raises at stope endlines and retreats to the timbered raise. This retreat sequence starts on the lowest sublevel and the opening cut is blasted into the slot raise. This retreating system is used in order to provide a good safe back over the miners working in any given sublevel. A crawler-mounted down-the-hole or conventional pneumatic drill puts the holes down from one sublevel to the next on any desired drill pattern. Gravity flow brings the broken ore down to the scram level. The illustration shows a large slusher mounted at the timbered raise and slushing from the pockets back to the raise chute. Development of the ore block in Fig. 2 is typical for a 5.5-m (18-ft) wide vein. Initial development is by a trackless heading driven in the footwall of the vein. Pockets are then put into and through the vein on 7.6-m (25-ft) centers. For ease in using trackless rubber tired mucking equipment, these pockets are crosscut to the vein at approximately 1.04 rad (60°) rather than per¬pendicular. A two-compartment timbered raise is driven down from one level to the next. Endline slot raises are again put through from level to level. Each of the sub¬level blocks selected in this example is 40-m (130 ft) long by 33.5 m (110 ft) high. Stoping starts at the slot raises and retreats back to the timbered raise. Long holes are drilled from one sub¬level down to the next by starting between the lowest pair of sublevels and then working up the ore block. Mucking out of the pocket drawpoints is accomplished by rubber-tired load-haul-dump (LHD) units which then dump into a centrally located orepass.

Jan 1, 1982

By Cuong V. Trang, Christopher A. Fleming

There has been greatly renewed interest throughout the gold mining industry over the last few years in cyanide recovery from gold and silver plant tailings. This interest has been spmred by two factors. Firstly, the increasing cost of destroying cyanide to levels that meet more and more stringent regulations for tailings impoundment or discharge to the environment. Secondly, many gold ore bodies suffer high cyanide consumption costs as a result of the presence of base metals (principally copper) that react with and consume cyanide during the gold leach. In most cases, the cost of recovering and recycling this cyanide from the tailings is signif­icantly lower than the cost of purchasing new cyanide. When this cost benefit is added to the lower costs of tailings detoxification, a compelling case can frequently be made for cyanide recove1y. A number of processes for recovering cyanide from gold plant barren solutions or pulps have been developed. Most of the processes require that the recoverable cyanide is first converted to the highly toxic hydrocyanic acid gas (HCN) and fears surrounding the handling of this compound have probably presented the great­est impediment to implementation of these processes.

Jan 1, 1998

By Alden B. Carpenter

The Dead Sea is the only relatively large body of surface water having bromide concentrations comparable to those in the brines of the Smackover Formation in Arkansas. The Dead Sea, like the Smackover brines, contains a great deal of calcium and is very low in sulfate and iodide. The relationships between potassium, divalent cations, and chloride indicate that over 20 percent of the bromide in the Dead Sea is derived from some source other than by the evaporation of sea water. A similar type of study of the Smackover brines in Arkansas indicates that approximately 60 percent of the bromide has been derived from some source other than sea water. Both the Dead Sea and Smackover brines are very low in iodide suggesting that the excess bromide has not been derived from organic matter. The origin of the excess bromide is probably the same for both of these brines, but the source of this bromide remains unknown.

Jan 1, 1978

By B. Das, D. D. Mishra, M. I. Ansari

Experiments conducted with a 75-mm-diem hydroryclone have shown it is possible to obtain an iron ore concentrate containing 60% Fe. 2.65% SiO2, and 4.0% Al2O3 from Barsua iron-ore washing-plant slimes. About 60% of the iron values are recovered from slimes containing approximately 49% Fe, 8.9% A1203 and 11% SiO2. The obtained products can be used in the sinter feed for iron and steel making. The key to achieving process success lies in establishing the correct cut point (d50) of the hydrocyclone through the manipulation of operating and design variables, such as spigot and vortex finder diameter, solids concentration and operating pressure. The cyclone efficiency is represented by the partition curve. The cyclone performance, evaluated mathematically from the reduced efficiency curve, is represented by the equation Y = 100[1 -exp(-axb)], where: •X = did 50, •Y = corrected of efficiency (%) and •a and b are constants.

Jan 1, 1994

By B. D. Hong

The Coal supply and demand review arrived too late to appear in the May issue. In 1991, the US coal industry produced 902 Mt (994 million st) of coal, according to preliminary data compiled by the Energy Information Administration (EIA) (Table 1). This amount was 32 Mt (35 million st) less than the record 933.5 Mt (1 billion st) produced in 1990. Most of the decrease in coal production occurred east of the Mississippi River. The substantial stock buildup that stimulated production in 1990 was not repeated in 1991. In addition, coal consumption by electric utilities and coke plants dropped by about 1.8 Mt and 4.5 Mt (2 million and 5 million st), respectively. Moreover, a higher level of coal exports failed to bolster the 1991 production. Production Surface mines accounted for about 59% of the 902 Mt (994 million st) of coal produced in 1991. Underground mines produced for 41%. More than one-fourth of the total coal output was from federal and Indian coal leases. Production of all ranks of coal, except subbituminous, declined in 1991. Bituminous coal is produced mainly in the East. It totaled an estimated 590 Mt (651 million st), a 6% decrease. Production of subbituminous coal, mostly in the West, rose 3% to a record 228 Mt (252 million st). Lignite output was from Texas and North Dakota. It decreased by 1 % to 79 Mt (87 million st). Anthracite production, all from eastern Pennsylvania, fell by 17% to 2.7 Mt (3 million st).

Jan 1, 1992

By J. M. Mutmansky, Y. J. Wang

The usual method employed to determine the direction of flow in a connecting airway in advance is to conduct a survey between the two existing nodes being connected by means of the new airway. It is then assumed that airflow will take place from the node with higher pressure to the node with lower pressure. The purpose of this paper is to examine this direction-of-flow problem -- using properties of the system characteristic curve for a two-terminal network -- and specify some conditions that ensure that flow will take from the higher-pressure node to the lower-pressure node. It is concluded that the only requirement is that all characteristic curves of fans in the network should be strictly decreasing. Five numerical examples illustrate some important properties of the direction-of-flow problem.

Jan 1, 1999

By P. D. Roa, O. Noirot, H. Owens, D. E. Walsh, J. R. Mokka

Over a period of eight years (1989 through 1997) the Mineral Industry Research Laboratory (MIRL) at the University of Alaska, Fairbanks, AK, has investigated the effects of temperature, residence time and particle size on the degree of low-rank coal (LRC) upgrading for a nonevaporative, hydrothermal drying process, also known as hot water drying (HWD). The following three Alaskan LRCs were used in this study: Usibelli coal (Seam No. 4) from the Nenana Coal Field, Little Tonzona coal from near McGrath, and Beluga coal from the Beluga-Yentna field on the west side of Cook Inlet. Replicated factorial tests were conducted. Experimental results indicate that most of the low-rank coal upgrading via HWD occurs rapidly within the first 10 to 20 min of residence time. Thereafter, the upgrading process slows down, and more than 100 min of additional residence time is required to produce percentage increases equivalent to what was achieved in the first 10 to 20 min. The properties of the hot water dried products were very sensitive to process temperature. Calorific value, carbon and oxygen contents, equilibrium moisture levels and Hardgrove gindabilities all appear to change linearly with temperature within the 275°C to 325°C (525°F to 61 5 OF) range.

Jan 1, 2000

By H. P. Ehrlinger

The Illinois State Geological Survey added Minerals Processing to its capabilities in 1966. Subsequently, the Minerals Processing Team and several Geologists collectively thought what neither would have thought of separately. The results included a number of projects. discussed here are: 1) The Utilization of the Fines from Kitty Litter as a Bonding Agent for Iron Ore Pellets; 2) Separating Plant Waste from a Commercial Silica Sand Producer into two marketable products, including a closely sized Silica Sand; 3) Marketable Feldspar Sand and Silica Sand from Dune Sands near the Kankakee River; and 4) A Method for Making Metallurgical Grade Fluorspar more desirable as Blast Furnace (Steel) Feed.

Jan 1, 2003

By P. B. Marchant

Bioleach testing and piloting was conducted on refractory gold sulphide concentrate from Austin Gold Venture in Nevada during 1988. The objective of the test programme was to determine the technical and commercial viability of bioleaching as a concentrate pretreatment step prior to gold extraction by cyanidation as an alternative to shipping concentrate to the smelter. Baseline CIL cyanidation typically resulted in < 20% gold extraction from the Austin concentrate. Bioleaching enhanced gold extraction up to 95% using conventional CIL techniques. This paper summarizes the bioleach test programme and results, including pilot plant studies onsite, and describes aspects of the commercial design for a possible retrofit of the Austin flotation tailings leach circuit to accommodate 20 tpd and 40 tpd capacity concentrate bioleach facilities.

Jan 1, 1989

By B. Govindarajan, T. C. Rao, M. Vanangamudi

The results obtained from hatch, flotation experiments on two coals were used to develop relationships between the recoveries of water and solids into the concentrate. When the recovery of non-ash material (RN) for any size fraction is plotted against a parameter X, defined as [x _ Amount of (W) recovered x Collector dosage (C)in the concentrate (cc)(kg/t)] a single "non-ash recovery curve" is obtained. Non-ash recovery curves thus obtained from different size fractions are used to get a unique "modified non-ash recovery curve" for a given coal, using the reduced efficiency curve concept. Similar observations have been noted in the case of recovery of ash material (RA) also. The amount of water recovered into the concentrate is plotted against a parameter "Water Recovery Factor (WRF)," defined as [WRF = T.F.(1+C)] where T is time (in sec) and F is frother dosage (in kg/t). Plotted in this form a characteristic water recovery curve, for a given coal is obtained that may be defined by the following equation: [W4=W_(1 pWRF-e] The above relationships have been verified against the experimental results obtained from three more coals and are found to give satisfactory predicted values.

Jan 1, 1992

"The profile of the respondents was that of a professional responsible for communications, community relations or corporate social responsibility programs. By targeting a consistent professional profile it was possible to ensure that most respondents had a similar degree of responsibility and/or decision-making ability related to the application of communication technologies. The research attempted to be inclusive and attention was paid to interview respondents from both genders and from indigenous groups.4, 5 A codebook of 80 themes was developed that highlighted recurring topics; for example, codes were created to identify risks, benefits, policies and learning process of using social media and included topics that were repeatedly emphasized by respondents (i.e. age, efficiency and language). Both manifest coding (i.e. coding of the visible content, such as specific words) and latent coding (i.e. coding of underlying meanings) was undertaken (Creswell [13]). The codes were used in the qualitative analysis using ATLAS.ti software and network views were created to illustrate the relationship between major themes of this research. There are limitations to this research, largely related to the changing nature of the Internet and communication technologies. There is also a geographic limitation to the application of this research because most of the data was gathered primarily from Canadian sources.6 Different levels of Internet penetration, or limited access to computers or mobile devices, and low literacy rates may reduce the applicability of the research in countries with low Internet penetration. A more systematic process to selecting candidates may have helped bring balance to the final list of interview respondents. For example, most of the corporate respondents were from medium and large companies and there was not adequate representation of small mineral exploration companies. The research did not probe into social media applications by different business groups within the mining sector, such as smelting and processing."

Jan 1, 2011

By Charles H. Jacoby

In 1960 the management of International Salt Company made the decision that all waste material from the Watkins Glen plant would be stored in underground salt cavities. The cavities which they contemplated using had been depleted of salt to a point where continued operation would have impaired their structural stability. Originally, the waste to be disposed of included vacuum pan bleeds, cinders, contaminated salt and sewage. It was proposed that this material, in a slurry form, was to be injected into the salt cavity through one of the abandoned wells. Brine, displaced from the cavity, was to be recovered from an adjacent well and subsequently injected into the Cherry Valley limestone, This Cherry Valley limestone which underlies a large portion of north central New York contains a connate water with high concentrations of ammonia, hydrogen sulphide and the chlorides of calcium, magnesium and sodium. A number of questions were immediately raised on various aspects of the technical feasibility of the project. Of major concern was the buildup of insolubles adjacent to the bore of the injection well. Another pertinent question was the carryover in the displaced liquor of finely-divided, mechanically-suspended particles. Criticism was leveled at the inclusion of sewage. Engineering studies were started by the company&apos;s Mineral & Mines Development Department and on April 3, 1961 a preliminary feasibility report was issued.

Jan 1, 1973