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By M Hassall, E Humphries

Personal risk assessment processes, also known as Take 5 or SLAM (Stop, Look, Assess, Manage) have been utilised in high hazard industries for over 30 years. However, there is limited research to understand if a personal risk assessment process improves safety and health outcomes. The author’s learnings during discussions with training participants in G3/G-MIRM (Risk Management Systems) and G2 (Risk Assessment for Supervisors) courses have provided them with significant insight into understanding if a personal risk assessment process is valued across a number of mining industry organisations. These discussions arise when the four-layer model is presented (Figure 1) where ‘Layer 4’ represents the last resort risk checking activity. In this context, training participants agree that this process has value. However, the consistent opinion is that the personal risk assessment process is not effective, that it is a ‘tick and flick exercise’ and is not achieving the desired outcomes.

Apr 16, 2024

By Lee G, Stitt P

The term "industrial minerals" covers a vary wide field. However in Australia this field tends to be characterised by the follow- ing: (1) Relatively low unit values. (2) Product specifications can be exacting compared to the metalliferous field. (3) Product specifications often have to be tailored to the needs of the end user. (4) Beneficiation plants tend to involve low capital outlays relative to the metalliferous field. However they often have to produce products'to a range of specifications. (5) Technology often influences what can be considered as reserves. (6) Because of low unit values industrial minerals tend to be relatively cost sensitive. This is particularly the case in regard to transport. (7) Location of deposits near to major urban areas (markets) dictates a greater pressure on ecological aspects. (8) In Australia there has been a tendency to vertical integration largely due to the restricted market available. Reserves assessment is shown to be critically dependent upon the above Factors.

Jan 1, 1979

?AARONSON, ALFRED E, .Vice-Pres, Mid-Co Petroleum Co , Mid-Co Bldg, Tulsa, Okla '18 ABADIE, EMILE R, Min Engr - Box 927, Porterville, Cal '76 ||ABADILLA, QuiRiCo A, Student, Colorado School of Mines, Golden, Colo. '20 ABBOTT, CLARENCE E, Gen'l Supt, Iron Mines Div, Tenn Coal, Iron & R R Co , 1405 Minnesota Ave, Bessemer, Ala. '04 ABBOTT, Louis BENJAMIN, Chief Engr, Consolidation Coal Co-, Elkhorn Div, Jenkins, Ky '15 ABBOTT, ROBERT R Met. Engr, Peerless Motor Car Co, Cleveland, Ohio. '12 ABE, ASAKA Fujita Min Co, Dojima, Osaka, Japan '14 ABEEL, GEORGE H, JR, Cons Min Engr, 633 St Paul Ave, Los Angeles, Cal '18 ABEL, ALBERT E, Mgr., Pittsburgh Dist, Valley Mould & Iron Corpn, 28th St & R R Ave, Pittsburgh, Pa '19 ABEL, W D, Min. Engr. 3897 Sacramento St , San Francisco, Cal. '19 ABELL, 0 J, Pres & Treas, Abell-Howe Co, 332 South Michigan Ave, Chicago, Ill '19 ABERNATHY, G E Willow Springs, Mo '20 ABEY, HIROSHI, Min & Met Engr, Arakawa Mine, Mitsubishi Co, Senpoku-gun, Akitaken, Japan '17

Jan 1, 1923

"Located on the Big Hole River, three miles from Divide, Montana, and 21.75 miles from Butte.Built in 1899.DAM: Rock-filled wooden crib, 512 ft. long and 57.5 feet high.SPILLWAY: Equipped with flash-boards 189 ft. long, 8 ft. 10 inches high.PENSTOCKS: Two 8 ft. diameter of steel, each branching and supplying two wheels.WHEELS: Four double horizontal, double discharge turbines, made by The James Leffel & Co., 1,000 h-p., 180 r. p. m., direct connected to generators, operating at 60 ft. head. GENERATORS: Four G. E., revolving field, 750 k-w., 180 r. p. m., 800 volt, 3 phase, 60 cycle alternators.EXCITERS: Two water-wheel driven exciters.TRANSFORMERS: 12 250-k-w. G. E. air blast, primary 15,000 volts, Secondary 800 volts.OPERATION: The plant operates essentially as a peak load plant. Except in flood period, there is insufficient water to operate at full load continually. In December its average output is about 500 k-w."

Jan 1, 1913

By Mark S. Stagg, Rolfe E. Otterness, Stephen A. Rholl

The Bureau of Mines fired seven test blasts in a 22-ft bench of limestone, screening the material to investigate the influence of explosive type and between row delays on fragmentation. Four 4-hole, single-row test shots were conducted evaluating dynamite, ANFO, emulsion, and a 60-40 emulsion-ANFO mix. Three 12-hole shots, with three rows of 4-holes, evaluated delays between rows of 24, 36, and 120 ms. Accurate in-hole electronic delays were used, firing within 1.5-ms of the nominal delay. All seven shots had a 6- to 7-ft burden, 10-ft spacing, with a hole diameter of 3-1/2 in fired on a 12-ms delay between holes in the row. The tests were instrumented with strain gauges grouted in the burden region examining stress wave and gas pressure effects. Fiber optic probes were used to confirm timing and to obtain detonation velocities.

Jan 1, 1989

By R. Steinitz, J. L. Orehotsky

An analysis was made of the influence of floating molten zone traversal on the purification of tungsten rods and the effect of this purification on the brittle -to-ductile transition temperature. The major impurities detectable in the original rods, i.e., molybdenum and oxygen, were reduced by zone refining from 30 to 3 Ppm. Changes in the trace amounts of other impurities could not be quantitatively determined since they were below the limit of detection. Tensile tests on 3-mil wires, drawn from these rods and recrystallized to 17,000 grains per sq mm, showed a transition temperature of 60°C for the zone Purified and 175°C for the original material. 1 HE brittle-to-ductile transition temperature is manifest mainly in the bee metals, iron,' molybdenum, chromium,3 tungsten,4 and niobium. Since many of these metals cannot be fabricated at room temperature in a commercially pure annealed poly-crystalline state, a considerable amount of work has been done on evaluating the influence of purification on the transition temperature in order to determine indirectly whether the fabricability of these metals can be improved by removing impurities. Smith and Rutherford8 showed that the transition temperature of iron is lowered significantly by zone purification. Working with chromium, Edwards, Nish, and wain7 found also that purification appreciably lowered the transition temperature. A similar evaluation has so far not been reported for tungsten, probably because of the difficulties in obtaining tungsten in a purified polycrystalline state. carlson8 and, more recently, schadler9 have purified tungsten rods by floating zone traversal, but after zone traversal, their rods were single crystals. To obtain high-purity polycrystalline tungsten, a technique has to be round for transforming the zone-purified single crystal material into a suitable polycrystalline structure. The purpose of this investigation was to obtain zone-purified tungsten in a suitable polycrystalline state and then to evaluate the effect of zone purification on the transition temperature.

Jan 1, 1962

By A. Shakoor, T. P. Gautam

The long-term slaking behavior of clay-bearing rocks in the field can be quite different than the short-term slaking behavior in the laboratory. To assess this difference, 12 replicate samples of each of 20 clay-bearing rocks, including 5 claystones, 5 mudstones, 5 siltstones, and 5 shales, were exposed to natural climatic conditions for 12 months. After each month of exposure, one replicate sample of each rock was removed from natural exposure and its grain size distribution was determined. Disintegration ratio (DR), defined as the ratio of the area under the grain size distribution curve of the slaked material for a given rock to the total area encompassing all grain size distribution curves of the tested samples, was used to quantify the amount of disintegration under natural climatic conditions. DR values of naturally exposed samples were then compared with the corresponding values of second-cycle slake durability index (Id2) determined in the laboratory. The relationship between Id2 and DR after, 1 month of exposure, exhibits an R2 value of 0.65 but the relationship deteriorates when DR values for 6 and 12 months of exposure are used, with R2 values of 0.63 and 0.25, respectively. However, when DR values for laboratory samples, computed from grain size distributions of slaked material after second cycle, are compared with DR values of naturally exposed samples for the three periods of exposure, the relationship improves with the corresponding R2 values being 0.61, 0.71, and 0.41. These results suggest that DR may be a better parameter for assessing the long-term disintegration behavior of clay-bearing rocks than Id2.

Jan 1, 2015

By C. K. Baxter, A. Odeshi, M. Mohammadi, E. D. Cyr

"Metal additive manufacturing has been a revolutionary step in designing new complex shapes with faster and cleaner building capacity in comparison to subtractive manufacturing processes. The ability to print lightweight alloys such as aluminum has made metal 3D printing attractive to different industrial sector from aerospace to energy and bioengineering. Of particular interest in this study, are the mechanical and microstructural properties of AlSi10Mg-200°C alloys. In this project, samples were 3D-printed in horizontal direction using an EOS M290 machine by Direct Metal Laser Sintering (DMLS) technique. Uniaxial tensile tests and high strain-rate impact e tests were carried out to investigate the mechanical properties and to study the repeatability of the process. A constitutive model is developed to predict the results of the mechanical tests at varying strain rates to better understand the high strain-rate behaviour of these alloys. Using optical microscopy, features such as scan tracks and melt pools within the AlSi10Mg-200°C samples were observed, and the microstructure was observed to be homogenous. The previously listed tests were compared to the results obtained for similar cast alloy counterpart, i.e. A360.0.INTRODUCTION A relatively new way of manufacturing parts is on the increase as rapid prototyping becomes increasingly popular throughout the world. Additive manufacturing is a process which, given a 3D model, uses metallic material in powder form to build a part layer by layer until completion (Gu, Meiners, Wissenbach, & Poprawe, 2013). This way of rapid prototyping is very attractive in comparison to other possibilities because of its ability to accommodate complex design and geometries. Conventional manufacturing methods have limitations when it comes to complex geometries, but additive manufacturing allows for increased design flexibilities while maintaining a light and stable final product (Murr et al., 2012). It also allows for a reasonable cost, especially for small batch sizes that would otherwise be extremely expensive."

Jan 1, 2018

By N. Ohmura, H. Saiki, T. Nagaoka

A microbiological flotation system that facilitates the removal of pyritic sulfurfrom coal is reported. The chemoautotrophic bacterium Thiobacillus ferrooxidans is utilized as a pyrite-selective biological surfactant that augments the flotation-mediated rejection of pyritic sulfur. The bacteria were produced by continuous cultivation in 50-L tanks, where it was possible to produce 3.8 x 1012 cells/day when the dilution rate was optimal and the pH was 1.6, a condition that prevented ferric precipitation. Prior to flotation, the bacteria were mixed with coal, which allowed cells to selectively adhere to the pyrite. The adherence of bacteria made the pyrite surfaces more hydrophilic, which facilitated separation by increasing the tendency of pyrite to sink, while the hydrophobic coal floated. When a synthetic coal/pyrite mixture (10.1 % sulfur) was subjected to microbial flotation, the level of sulfur rejection was stable during the entire period of operation (i.e., one hour). In the absence of bacteria, flotation reduced the sulfur content of a coal/pyrite mixture to within the range of 6.7% to 7.4%, with the combustible recoveries ranging from 93.9% to 99.0%. In the presence of bacteria, the sulfur content of cleaned coal was in the range of 1.3% to 1.4%, with the combustible recoveries ranging from 80.8% to 86.3%. Bacterial activity accounted for 57% of the total sulfur rejected through flotation. The desulfurization capacity of the system was further confirmed by testing it with run-of-mine coal. The pyritic sulfur content was reduced from 2.9% in the feed coal to around 1.2% in the froth product, while 80% of the combustible recovery was retained. Microbial flotation removed 60% of the pyritic sulfur from the feed. These findings demonstrated that micro¬bial flotation could be utilized for the removal of pyritic sulfur from coal in a semicontinuous system.

Jan 1, 2000

By G. F. Sager, B. J. Nelson

Magnesium alloy forgings offer higher and more uniform mechanical properties than heat treated magnesium alloy castings and are used principally for light weight parts that may be stressed in fatigue or subject to impact loads, or where space limitations do not permit the use of the necessarily bulkier magnesium alloy castings. Compositions and Properties of Magnesium Forging Alloys The nominal compositions and properties of the principal commercial magnesium forging alloys are given in Table 1. The two that are most commonly used are AMC58S and AM65S. The former has the highest strength and is, therefore, employed for heavy duty applications. However, the forging of this alloy and of AMC57S ordinarily requires the use of slow speed hydraulic presses. The forging alloy AM65S combines good, although somewhat lower, mechanical properties with excellent forging characteristics. This alloy can be forged on hammers, upsetters, rolls, and mechanical presses, as well as on hydraulic presses. These characteristics make it particularly useful for many applications. Manganese Segregation in AM65S The alloy AM65S contains 5 pct tin, 3 1/2 pct aluminum, and a small amount of manganese, the manganese being added to insure satisfactory resistance to corrosion. Some manganese segregation was occasionally encountered in this alloy. In many applications, such segregation is of no particular consequence, but when forgings are acid dipped prior to the application of various chemical coatings, the segregated particles of manganese constituent may stand in relief. The resultant surface roughening may impair appearance and is undesirable in forgings employed in the textile industry, where very smooth surfaces are essential to avoid the catching and breaking of fine threads. An effort was therefore made to determine the cause of manganese segregation in this alloy and find a method for its elimination. Radiographic examination of a series of cast ingot sections and extruded samples of the AM65S type having manganese contents ranging from zero to about 0.7 pct revealed no appreciable manganese segregation when the manganese was less than about 0.4 pct. A small amount was observed with a manganese content of 0.46 pct and considerable with manganese contents of 0.6 pct or higher. Attempts to eliminate the manganese segregation by variations in the melting and casting conditions were unsuccessful and it soon became apparent that the trouble was attributable to the fact that the manganese contents that caused segregation were in excess of the amounts that would remain in solution at temperatures encountered in the melting and/or casting operations prior to the complete solidification of the ingots. A knowledge of the amounts of manganese that would be soluble in molten alloys of the AM65S type at various temperatures was obviously desirable. The liquid solubilities of manganese in binary Mg-Mn alloys and in Mg-Al-Mn and Mg-Zn-Mn alloys have been determined by Tiner and others,* but no data on the solubility of manganese in molten magnesium alloys containing tin and aluminum are given in the literature. For this reason, an investigation to determine the solubility of manganese in a magnesium alloy containing 5 pct tin and 3.5 pct aluminum (AM65S type) was undertaken.

Jan 1, 1950

By Mary Kama, Kaul Gena

"Magnetic skarn ore (MSO) is one of the major copper bearing ore extracted by the Ok Tedi Copper Mine in Papua New Guinea. Copper minerals are recovered by flotation while the iron not associated with copper are discarded as tailings. The objective of this investigation was to determine the iron ore reduction kinetics for the Ok Tedi MSO and ascertain if it can be processed to produce sponge iron for a mini steel plant in Papua New Guinea.SEM-EDAX analyses of the Ok Tedi MSO indicated 10.1% C, 30% O, 0.6% Mg, 1.1% Si, 21.1% S, 0.8% Ca and 36.2% Fe. Most of the iron is in sulfide form. Both naturally occurring and roasted sinters of Ok Tedi MSO samples of different particle sizes were reduced by charcoal carbon at three different temperatures and seven different reduction times. Analyses of the reduced products indicated a metallic iron content of more than 65 wt.% on average.Results showed that there was no significant difference in reduction between fluxed and control materials. Only a slight increase in kinetics with reduced particle size, hence the reaction rate constant (K) did not vary much within the temperatures investigated. Reaction kinetics increases with increasing reduction time at 900°C. Therefore, more iron reduction is observed with particles of 106 µm size. In addition, the results also confirmed that the reduction energy used was higher at 800°C and lower at 1,000°C. In summary, iron reduction can be improved but close monitoring of temperature and reduction times are essential to determine the reaction kinetics of the Ok Tedi MSO. INTRODUCTIONMost mines in Papua New Guinea mine large quantities of iron in the form of sulfides, carbonates, oxides and sulfates that are eventually disposed off as tailings. The Ok Tedi magnetic skarn ore (MSO) contains a significant amount of hematite, magnetite, siderite, ilmenite and pyrite that can be a source of iron. The mineralogical compositions of Ok Tedi MSO reveals more iron minerals than copper, yet all iron containing minerals are collected in the tails from the main copper production stream. Therefore, this study is conducted to investigate the possibility of producing direct reduced iron (DRI), or sponge iron, from the MSO. Such sponge iron may serve as the feedstock for electric arc/induction furnace based mini steel plants in Papua New Guinea."

Jan 1, 2016

The Douglas mineral sands deposit is located 75 km north of Hamilton in western Victoria, Australia. The deposit was discovered in the late 1990s by Basin Minerals Limited and purchased by Iluka Resources Ltd in 2002. Prestrip mining commenced in December 2004, with ore mining commencing in August 2005. Douglas is the first of several projects being implemented by Iluka Resources Ltd in the Murray Basin, with all heavy mineral concentrate planned to be treated at the companyÆs mineral separation plant in Hamilton. The mineralisation at Douglas lies along the south-eastern margin of the Murray Basin. It is hosted by marine Pliocene Loxton-Parilla Sands and consists of three discrete deposits; Bondi Main, Bondi West and Bondi East. The mineralised sediments dip at a low angle to the west and lie unconformably beneath a veneer of top soil (50 - 100 mm) and subsoil (up to 600 mm) and between three to 12 m of overburden. The barren overburden, characterised by high clay content (<53 ¦m material) generally provides a stark contrast to the sandy ore, which is between 1 m (parts of Bondi East) and 15 m (Bondi Main) thick. Both overburden and ore are removed utilising standard dry mining techniques using both scraper-bulldozer and truck-excavator combinations. All ore is delivered to run-of-mine (ROM) stockpiles before being reclaimed by front end loader into the processing plant. Grade control at Douglas is based around visual dig control aided by high precision survey, which enables accurate reference to the resource drilling and block model data. Vertical excavator faces, inclined scraper faces plus regular test pits and trenches provide opportunities for in pit mapping and sampling. This data is used to refine not only the resource estimate for Douglas but also the drilling, interpretation and resource modelling methods used for similar deposits across the basin. The foundations of the geological reconciliation process at Douglas are the in pit surveys undertaken each time the mining fleet has completed building a ROM stockpile. These surveys and the subsequent interrogation of resource models enable robust estimations of the heavy mineral content of the stockpiles, which are subsequently used to predict the daily input grades into the plant during their reclamation. Reconciliation of this daily data is then undertaken at the end of each month against metallurgical production data and plant balances. Learnings and results from the reconciliation process are used to further refine subsequent resource models and modelling methods.

Jan 1, 2009

By Lee B. Paterson

Mining Engineers throughout the world have an appreciation for the importance of the role of the Blasthole Drill in one of the primary elements of the mining process. This paper highlights areas that could influence the overall cost effectiveness of that process of breaking rock and the associated downstream mining and ore recovery activities - and how the drill is designed and configured to meet that challenge. Factors influencing drill productivity and overall cost effectiveness are addressed, particularly regarding the benefits which are not necessarily reflected in the drilling process and therefore more difficult to quantify. This paper explores some of the downstream benefits of what we do to the drilling process.

Jan 1, 2000

By M. R. Shahan, C. E. Seaman, S. E. Mischler, T. W. Beck

"Float coal dust is produced by various mining methods, carried by ventilating air and deposited on the floor, roof, and ribs of mine airways. If deposited float dust is re-entrained during a methane explosion, without sufficient inert rock dust quantities, this float coal dust can propagate an explosion throughout mining entries. Consequently, controlling float coal dust is of critical interest to mining operations. Rock dusting—adding inert material to airway surfaces—is the main control technique currently utilized by the coal mining industry to reduce the float coal dust explosion hazard. To assist the industry in reducing this hazard, NIOSH’s Pittsburgh Mining Research Division (PMRD) initiated a project to investigate methods and technologies to reduce float coal dust in underground coal mines through prevention, capture, and suppression prior to deposition. Field characterization studies were performed to determine quantitatively the sources, types, and amounts of dust produced during various coal mining processes. The operations chosen for study were a continuous miner section, a longwall section, and a coal handling facility. For each of these operations, the primary dust sources were confirmed to be the continuous mining machine, longwall shearer, and conveyor belt transfer points, respectively. Respirable and total airborne float dust samples were collected and analyzed for each operation and the ratio of total airborne float coal dust to respirable dust was calculated. During the continuous mining process, the ratio of total airborne float coal dust to respirable dust ranged from 10.3 to 13.8. The ratios measured on the longwall face were between 18.5 and 21.5. The total airborne float coal dust to respirable dust ratio observed during belt transport ranged between 7.5 and 21.8. INTRODUCTION During the extraction of coal, various dusts are produced that can be classified into two categories of interest: respirable and float coal dust. Respirable dust (diameter 10.0 = µm) has the capability to penetrate beyond the terminal bronchioles and become deeply embedded into the human lungs, leading towards respiratory disease complications (Potts et al., 1990). Float coal dust is defined as coal dust having a diameter less than 74 µm (Harris et al., 2009). During the mining process, float coal dust can be deposited on the floor, roof, and ribs of mine airways. With the proper mixture of methane gas (between 5% and 15%) and an ignition source, a methane gas explosion can occur (Barker and Humphreys, 1996). The resulting explosion can develop a pressure wave which lifts the deposited float coal dust material into the air, propagating a disastrous coal mine explosion (Harris et al., 2010). The current technique to safeguard against the propagation of a coal mine explosion is to treat the mine airways with incombustible rock dust (typically limestone), maintaining an 80% total incombustible content of the deposited material. NIOSH initiated a project to investigate control technologies that could be aimed at reducing the amount of float coal dust allowed to deposit in the mine airways using three approaches: prevention, capture, and suppression prior to deposition. Reducing the float coal dust being generated during the mining processes and capturing or suppressing airborne float coal dust prior to deposition would in turn increase the efficiency of the applied rock dust used to prevent mine explosions."

Jan 1, 2017

By Heihachi Kamura

The reduction of iron oxide, such as Fe203, to iron in the blast furnace is performed principally by carbon monoxide, but partly by solid carbon by the two following reactions: Fe2O3 + 3CO = 2Fe 4- 3CO2 2Fe2O3 + 3C = 4Fe + 3CO2 It is known that reduction by carbon monoxide takes place at a much lower temperature than reduction by carbon and that the reducing action of earbon monoxide progresses to the inside of the ore in lump form by gaseous diffusion. In the reduction by carbon, however, the reducing action takes place only at the contact surface of the ore and carbon, so that the more complete the reduction by carbon monoxide, the more efficient is the operation. However, the reduction does not progress as shown in the foregoing equation, because carbon dioxide acts oxidizingly on the metallic iron to such a degree that when the gaseous composition reaches a certain proportion of carbon dioxide to carbon monoxide, the reduction practically stops, even though the gases are moving slowly through the heated ore. The proportions of carbon monoxide and carbon dioxide in the gas depend on the kind of iron oxide and the temperature of reduction, also on the duration of time in the reaction. The longer the time of the reaction, the higher the proportion of carbon dioxide obtained in the gas, but finally it will attain to a condition of equilibrium. Chemical Equilibrium of Iron, Oxygen, and Carbon Investigations concerning the equilibrium between iron, oxygen, and carbon have been carried out by E. Baur and A. Glaessner,1 who deter-

Jan 1, 1925

By B. Seegmiller, W. A. Hustrulid, O. Stephansson

Introduction The concept of in-the-wall haulage was formally introduced to the mining community in the early 1970s. The basic concept is simple. Rather than the traditional in-the-pit haulage/access ramp, single or twin ramps driven in the pit wall are used with crosscut connections on the working levels. The in-the-wall pit haulage system could be considered as part of the opening of a new mine, a way of completing to final depth an existing pit, or part of the transition from open-pit to underground mining with the ramp remaining as part of the access. This paper is intended to acquaint the reader, particularly mine planning engineers, with some of the possibilities presented by the concept. Nordstrom (1983) said, "In many countries, mining engineers are either open-pit or underground specialists and, hence, often don't consider the possibilities for combining the best aspects of each into a hybrid mining system." This is believed to be true. There are few "hybrid mines" employing, for example, open-pit methods with underground haulage systems. One reason is the variance with the experience of the mining company and the engineers involved. However, with today's low metal prices, innovative and unconventional thinking may be just what a property requires to remain in production. All the possible situations where an in-the-wall ramp might be considered cannot be presented here. Rather, a few examples will be discussed for illustration purposes. Potential savings in stripping costs A major cost in open-pit mining is associated with the removal of overburden and waste rock. Although much of this stripping cannot be avoided, the volume (V) associated with the presence of an in-pit haul road can be quite large, V 100 (pit depth) a (road width) 2 x road grade (%) As an example, consider a vertical ore body 400 m (1312 ft) long, 60 m (197 ft) thick, extending to a depth of 160 m (525 ft). Cross sections are shown in Fig. 1. Other data are: ore tonnage = 11 Mt (12 million st), ore density = 2.9 t/m3 (0.09 st per cu ft), waste rock density = 2.7 t/m3 (0.08 st per cu ft), road width = 20 m (66 ft), road gradient = 10%, and level interval = 10 m (33 ft). The evaluation has been made assuming that the pit slope (0) is constant around the pit except on the side with the haul road. On this side, the slope is decreased to allow for the road. Slope angles of 30° to 60° have been selected to indicate the dependence of stripping volume (V) on slope angle. The stripping at the ends have been calculated assuming representation by cones. The volume associated with the haul road is 2.6 Mm3 (92.8 million cu ft). A comparison of the strip volumes involved is shown in Table 1. As can be seen, the percent savings in strip volume varies from about 4% to 17% and depends greatly on the slope angle. It is perhaps more interesting to examine the savings in stripping costs that could then be used to cover the cost of the access ramp and the crosscuts for stripping costs of $3/m3 ($0.08 per cu ft). The cost with the road volume is 2.6 million x 3 = $7.8 million.

Jan 1, 1988

By Robert C. Edmiston

The measurement of temperatures in exploratory drill holes yields information on regional and local geothermal gradients and mean surface temperatures. This data may be beneficial to the mining geologist interested in the occurrence and oxidation of metallic sulfide minerals or to the mining engineer interested in subsurface temperatures or water movement. Additionally, the estimation of mean surface temperatures by the projection of thermal gradients to the ground surface provides data on which to evaluate infrared imagery as an exploration tool. Temperature measurements in 30 diamond drill holes in the Silver Bell mining district of Arizona show that a regional geothermal gradient of 2.45°C per 100 m increases to 3.03°C per 100 m near the ground surface. It is also found that an anomaly in the mean annual surface temperature of 0.5°C is associated with the shallow occurrence of disseminated sulfides. Changes in the geothermal gradient due to topography, water circulation, and formation changes also occur.

Jan 1, 1976

By N. P. Gallagher

The U.S. Bureau of Mines investigated a hydrothermal sulfidation (HTS) and chloride leaching procedure to recover lead and silver from plumbojarosite. The plumbojarosite, generated during pressure-oxidative leaching of zinc concentrate in a commercial operation, contained, in percent, 25.4 Pb, 18.8 Fe, 1.4 Zn, 29.5 sulfate (S042.), 3.2 elemental sulfur (SO), 13.4 total S, and 7.4 tr oz/st Ag. The HTS was conducted in an autoclave and converted the plumbojarosite into amorphous iron oxide and liberated the Pb and Ag as sulfides (S2) The best operating conditions were 50 g of plumbojarosite and 7 pct So (2.04 g So added) in 500 mL of 2.0-mol/L NH3 at 150° C for 1 h. The sulfidation product contained, in percent, 37 Pb, 28 Fe, 2 S042., 7 total S, and 11.2 tr oz/st Ag. Flotation of the sulfidation product was conducted at pH 3 with 15-pct pulp density and 2.4lb/st of isopropyl xanthate as collector. The flotation concentrate contained 43 pct Pb, 25 pct Fe, and 13 tr oz/st Ag. FeCl3 leaching of the sulfidation product in a resin kettle extracted 98 pct Pb and 97 pct Ag from the sulfidation product, and HCI-Oz leaching in a shaker glass bottle extracted 99 pct Pb and 98 pct Ag. Both leach residues met the Environmental Protection Agency's Toxicity Characteristic Leaching Procedure (EPA-TCLP) standards for nonhazardous wastes. The HTS and chloride leaching procedure is a potentially effective method for detoxification of plumbojarosite-bearing residue from the zinc pressure-leaching circuit.

Jan 1, 1989

By Kenneth Cottingham

The year 1941 brought a considerable increase in drilling activity in Ohio. In 1940, the total of wells completed was 1228, whereas in 1941 it was 1561, or an increase of approximately 27 per cent. Of the 1561 wells, 333 were oil wells, 701 gas wells and 527 dry holes. The completions for the past five years are shown in Table I; the number of productive oil or gas wells for the year 1941, arranged by producing formations, are given in Table 2; the number of wells, oil, gas or dry, completed in each county in 1941, in Table 3, and wells by sands in Table 4. The production of oil increased by about 500,000 bbl. in 1941, nearly all of the increase being in Pennsylvania grade. In 1940, the total production was 3,052,000 bbl. from 25,500 wells. In 1941 the total production was 3,547,534 bbl. from 24,801 wells. The increase in production is chiefly due to the Clinton pool in Clayton township, Perry County, which is briefly men- tioned later in this review. The number of oil wells, by grades of oil, and the total quantities produced are shown in Table ;. Annual oil production for five years is listed in Table 6. The average prices paid for crude in 1941 were slightly higher than in 1940 These 1941 prices, with dates of change, are shown in Table 7. The maximum production of natural gas in Ohio was in the year 1915, when the total was just under 80 billion cu. ft. The maximum consumption occurred in the year 1916, and was approximately 169 billion cu. ft. For the past several years, Ohio has ranged from eighth to eleventh, in order of gas production for the different states, being in tenth place in 1940. In the matter of consumption, however, Ohio remains in fifth place. Preliminary figures for 1940 indicate that production of natural gas increased II per

Jan 1, 1942