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By A. W. Hesse

The trend of American construction toward permanence and longevity is noticeable in the more recent coal plant installations; also, the policy of many coal operators has changed from seeking to obtain the desired tonnage from several plants of small capacities to designing a single installation to produce the whole amount. The mine openings and plant of the Nemacolin mine of The Buckeye Coal Co., are located on the Greene County side of the Monongahela River, 73 miles south of Pittsburgh. The map (Fig. 1) shows the general location, also the Lambert syncline. This syncline partly accounts for the selection of the site; the other reasons were the possibilities of rail and river shipments. Two shafts and a slope were sunk to the coal, work starting on all three in the second half of the year 1917. The air shaft was sunk to a depth of 243 ft. and completed in February, 1918; the slope was completed in November, and the skip shaft was stopped at a depth of 21 ft. below the coal, then the concrete lining was stopped 8 ft. above the coal August, 1918. Work on the skip shaft was resumed May, 1923, and sunk to a depth of 60 ft. below the coal by September, but it was not until May, 1924, that the lining was poured. Max. Min. WateR Water during after Sinking, Crouting, Depth or Cu. Yd. Cu. YD. Gal. per GaL. per Opening Size Length Excav. Conc. Min. Min. Air shaft..... 20 ft. diam. 251 ft. 8 in. 4788 1902 475 2 Slope........ 11 it. by 19 it. 1 in. 857 ft. 11500 5200 50 2 Skip shaft.... 32 ft. 8 in. by 12 it. 8 in. 296 ft. 6042 2293 300 3 Fig. 2 shows sections of these openings. The surface plan shows the relative positions of the openings. Coal was first hoisted from the air shaft May 7, 1918, and entry driving was continued; but the coal had to be stored, as the railroad had not been extended from Crucible to Nemacolin. As soon as the slope was driven to the coal and the linings concreted, the entries were driven along projected lines to connect with those advancing from the air shaft, and skip shaft. As soon as the underground connections had been made between the two shafts, the hoisting by buckets was inaugurated at the skip shaft and construction on the fan at the air shaft started.

Jan 1, 1927

By Hinde J. S

The Fitzpatrick Orebody is a three and a half million tonne fault bounded portion of the early-middle Proterozoic Broken Hill Orebody.The Fitzpatrick Orebody was discovered in 1975 by diamond drilling from the 36 Level Exploratory Drive. This followed persistent geologically oriented exploration and a courageous decision by mine management to deepen No. 3 Shaft to provide exploratory development across the Globe Vauxhall Shear and allow effective drilling of the target area.The stratiforrn oebody consists of three lenses, the lead rich No. 2 and 3 Lenses and smaller zinc lode. They represent a down plunge extension and boudinaged, fault-bounded portion of the main Broken Hill ore lenses. The surrounding host rocks are extensively faulted and jointed due to complex history of movement on the bounding shears. This results in extremely poor ground conditions during mining. The ore was delineated for mining development using three dimensional fan drilling from exploratory and hanging wall drive drill sites.

Jan 1, 1990

By P. D. Kovscek

The Bureau of Mines equipped a longwall shearer with a water-jet-assisted mechanical cutting system to determine the effect of the high-pressure sprays on machine performance, airborne dust levels generated, and size of particles formed. Testing was conducted on a simu-lated coal block (coalcrete) while cutting at a water pressure of 190 psi and at water-jet-assisted pressures ranging from 1,000 to 6,000 psi. The shearer motor energy used by the shearer while cutting coalcrete was determined for each cutting condition. Results indicate that the motor energy required by the shearer for each ton of material cut did not vary significantly as the water pressure was increased. Results of dust measurements recorded near the cutting drum indicate that at water pressures above 3,000 psi, dust reductions up to 85 pct were achieved com-pared with cutting with conventional sprays. Analysis of the particle size distribution during cutting showed that increasing the water pres-sure resulted in increased median particle size. In light of these results, various mechanisms for water-jet-assisted cutting are discussed.

Jan 1, 1986

By C. C. Clark

The national Institute for Occupational Safety and Health (NIOSH), Office of Mine Safety and Health, Ground Control Engineering Branch is investigating the use of shotcrete in weak rock mass mines with the objective of reducing fatalities and injuries resulting from rock fall accidents. When shotcrete is used as part of a multi-element ground support system, it is necessary to know when the material has developed a compressive strength of approximately 1 MPa (145 psi), the early threshold for safe re-entry of miner and machine into shotcreted mine workings at this value, the material has developed enough strength to be self-supporting and to allow for installation of other ground support components that require drilling of the shotcrete layer without degradation. NIOSH researchers have developed methods and portable test equipment to measure shotcrete strength on site in the first six hours after application using a partial beam test standard, a STM C 116-90 (1990). These advances were demonstrated in tests with five commercially available shotcrete mixes, sprayed as dry shotcrete using field expedient methods and equipment. The strength values from these tests allowed for real-time identification of the early strength threshold and were consistent with strengths reported using laboratory-type equipment. Measuring the early strength properties of shotcrete directly at the mine site can improve mine safety by identifying appropriate re-entry times and providing a convenient means of conducting on site quality control during shotcrete applications.

Jan 1, 2011

By D. Warkentin, M. Tomlinson, A. Nacu

"This study focuses on comparing the “blinding” effectiveness of various organic additives as part of a general review of using organic additives to mitigate preg-robbing and improve gold recovery during cyanide leaching of carbonaceous gold ores. The laboratory results compared the impact of various organic additives on the prevention of gold adsorption using both activated carbon and ore samples containing carbonaceous matter, generally referred to as total carbonaceous matter (TCM). Although several organic reagents were found to have a significant blinding effect on activated carbon, these results did not translate into increased gold leaching efficiencies with actual ore samples.INTRODUCTION Preg-robbing refractory ores, which contain carbonaceous matter, have been problematic in cyanide plants for many years. In industry, preg-robbing by carbonaceous matter may be mitigated using various techniques including flotation, to remove the carbonaceous matter; addition of blinding agents, to passivate the carbon surface and prevent adsorption of gold cyanide complex; roasting, to partially or fully destroy the carbonaceous matter; chemical oxidation (Abotsi & Osseo-Asare, 1986); pressure oxidation; or microbial deactivation (Dunne, Staunton, & Afewu, 2013). Carbon-in-Leach and Resin-in-Leach are also used as process alternatives in which activated carbon or resins are used to compete with the carbonaceous matter for gold cyanide adsorption. For highly preg-robbing ores, thiosulphate leaching may be considered as an alternative to conventional cyanide chemistry (Baláz, et al., 2000). Natural carbonaceous matter is composed mainly of hydrocarbons, organic acids, and graphitic/amorphous elemental carbon. Natural elemental carbon that behaves as activated carbon is considered to be the main reason for preg-robbing, although the true nature of this phenomenon is not fully understood (Tan, Feng, Lukey, & van Deventer, 2005). In general, the hydrocarbon fraction is assumed to have negligible interaction with gold, while organic acids, such as humic acids, may form complexes with gold, inhibiting its extraction in cyanide solutions. Organic reagents have been used for blinding carbonaceous matter in gold ores prior to cyanidation. Abotsi & Osseo-Asare (1987) tested various organic compounds, with two possible effects in mind. First, the organic compounds were expected to adsorb onto the surface of the solids as anionic/cationic species; the resulting electrostatic interactions between the Au(CN)2- anion and the negatively/positively charged surface would inhibit/enhance the gold uptake. Second, the organic reagents were expected to coat the surface of the carbonaceous component in the ore, limiting the available adsorption sites and reducing gold losses. Optimal reagent dosage depends on ore mineralogy, maturity and surface area of the carbo"

Jan 1, 2019

By J. R. Thoenen

The objective of this circular is threefold: First, to describe briefly the various modes of occurrence of sand and gravel. Second, to emphasize the need for adequate prospect¬ ing and exploration prior to the expenditure of funds for development and plant construction. Third, to describe various methods of prospecting and exploration and their application to the several modes of occurrence of sand and gravel. Sand and gravel may be found in nearly every county or township in the United States. Commercial deposits, however, are by no means so widespread. The introduction of concrete and the consequent impetus given to consumption of gravel and sand aggregates has seriously depleted deposits in some congested areas. Increasing re¬ strictions in the specification requirements for marketing sands and gravels have further narrowed the economic areas. On the other hand, the combination of increasing transportation costs and decreasing selling prices has forced the decentralization of gravel production. As a result smaller, intermittently operated, and widely scattered portable or semiportable unit plants have been encroaching on the territory formerly supplied from centrally located pits with large tr.atment plants. Generally speaking, the small portable plant can not produce any great variety of ma¬ terials to as exacting specifications nor as economically as the large permanent plant.

Dec 1, 1932

By Racquel Nottingham, Marc Mastrantuono, Freddy Lopez

In the digital era that we live in, connectivity and mobility play a decisive role in our daily activities. In this context, insufficient telecommunication infrastructure gives origin to the so called “dead zones” (areas with poor or non-existing cell phone reception), which hinder economic development in small communities and in many rural areas, since they remain invisible to a larger marketplace due to the inadequate cellular coverage. Several countries are working to identify these dead zones and with this information, establish strategies together with mobile providers in order to improve the existing mobile networks by expanding the coverage areas. As a result of the analysis, it is expected that thousands of telecommunication towers for 4G and 5G networks will be erected in the coming years. Telecommunication towers are slender elements that require large shallow foundations to verify primarily the safety against overturning. However in remote areas, it can be advisable to consider alternatives to the conventional shallow foundations, in order to optimize the use of materials (concrete, reinforcement steel, etc.). The following article presents an overview of intervention projects, where telecommunication towers were founded on micropiles. The micropiled foundations allowed to reduce considerably the dimensions of the originally planned shallow foundations. Furthermore, the use of micropiles allows for additional optimization, in combination with pre-fabricated foundations (i.e., precast concrete elements, etc.). This paper will focus mainly on the design and the execution of the micropiled foundation.

Sep 8, 2021

By D. M. Ziebarth

"Normal sequence of operations to develop a new underground mine is: construct and outfit headframe for shaft sinking, sink shaft, construct and outfit shaft sump, develop shaft station and shaft bottom steel. In an effort to reduce the overall project schedule, an innovative plan was developed to install service shaft steel below station level, prior to the completion of the shaft sinking effort. To better evaluate the construction of this substantial steel tower, a preliminary constructability evaluation was conducted, including the best method of supporting the main tower; from the bottom, middle or top of the tower. The design criteria and plan involved: •,Develop access from existing service shaft, 25 km underground •,Use of materials to maximize strength while minimizing corrosion exposure •,Estimating the location of the shaft centerline •,Over excavating the station and shaft below station to allow future shaft steel repositioning •,Designing the tower steel for modular installation •,Incorporate commissioning considerations, such as suspension of the cheese weights during steel installation •,Provision for covered stair access to shaft bottom •,Enclosed shaft envelope with open panels to allow visibility and ventilation The design approach was innovative and this paper will describe the design principles and challenges encountered while providing insight into the solutions developed and implemented; especially the modular construction and installation methodology. INTRODUCTION In an effort to expand the mining operations, a plan was developed to construct a new service shaft closer to the new mine workings and convert the existing service shaft to a production shaft at the Potash Corporation of Saskatchewan Inc. (PCS now Nutrien) Rocanville Potash mine in southeast Saskatchewan. The new service shaft required a multi roped friction hoist operating a cage / counterweight, rope guided conveyances to support the operational requirements. The service shaft is an integral part of the mining infrastructure as it provides access to the underground workings, transport personnel and materials underground for operation and maintenance of the mine. The shaft also allows for services required for safe operation of the underground mine, power supply, controls, communication, dewatering and fresh water supply."

Jan 1, 2019

By Atmesh K. Sinha

Death and injury rates in mining have been the highest of any industry; the injury severity rates in mines have also been the highest. Tentative figures reveal that 1971 was the best year yet for the bituminous coal indus¬try with the lowest number of fatal accidents. Nevertheless, 177 men were killed at a frequency rate of 0.76 per million man hours. Coal mine accident statistics have been described at various times as unacceptable and intolerable and recently resulted in the enactment of massive legislation. The 1969 Coal Mine Health and Safety Act was passed by Congress "to improve the working conditions and practices in coal mines and to prevent accidents and occupational diseases originating in the coal mining industry." The prevention of accidents in coal mines is the statutory responsibility of mining engineers and the moral responsibility of all who have any part to play in mining coal. The questions to be answered by these men may well be - What are the reasons for the high accident rate? Are the mines as safe as they can be? Are our mines sufficiently manned to insure proper supervision of production cycles and safety requirements? Are our plans for safety committees, safety campaigns, and safety inspections sufficient? Are they producing the results intended? If not, how can these be reorganized to have a greater impact? What is the real cost of an accident? What are the immediate and long-term plans for further improvements? These questions are in no way exhaustive, and the answers are not easy. There are many aspects of mine safety, and any one of them can be the subject of a lengthy discussion. The aim of this paper, therefore, is to limit the analysis to documented fatal electrical accidents in mines so that meaningful conclusions for improvement in electrical accident-experience statistics can be made. In doing this, it is also the intention to emphasize areas of missing information and propose a few suggestions for better hazard analysis.

Jan 1, 1972

By G. Eaton

Safety management is an important mortal and business function in construction. Contractors have traditionally tracked and reported lagging indicators, e.g. fatalities and lost-time accident rates, to measure their safety performance and stay in compliance with relevant regulations. Over the last several decades, contractors became more proactive in their approach to safety and developed programs that track leading indicators, e.g. safety audits and safety climate. In particular, safety climate measures workers? perception of safety management and its effectiveness in the workplace. Past literatures indicated that, while some industrial construction contractors ignore safety climate measures, others are limited by a lack of a formal means to measure safety climate. In this study, a survey approach is used to measure an industrial construction contractor?s safety climate through three key areas: management commitment, job control, and general safety climate. As a pilot study, a total of 214 individuals at a fabrication facility participated in the survey to verify the validity and effectiveness of the proposed survey approach. This survey study also confirms that job control and management commitment have a positive correlation and that worker demographics have an effect on respondents? perceptions of management commitment.

Aug 1, 2013

By N J. Welham

The Intec Copper Process is an environmentally advantageous hydrometallurgical process for the production of high purity copper and associated precious metals from copper sulphide concentrates. The process uses a mixed chloride-bromide lixiviant in an elegant cyclic circuit to leach the copper into solution, rejecting the iron as stable hematite rather than as unstable jarosite. After purification of the pregnant liquor, copper is electrowon at London Metal Exchange Grade A purity, while the anodic energy of the cell is stored as the soluble species, HalexTM, for recycle as the regenerated lixiviant to the leach circuit. Both gold and silver are leached and recovered directly from this cyclic process without resorting to cyanide leaching of the residue. No liquid emissions result from the Intec Copper Process, whilst spent air and water vapour from the leach are released to the atmosphere. The sulphur in the minerals reports to the solids residue in elemental form, without the need for expensive handling of voluminous gaseous sulphur dioxide streams found in smelters. A significant difference to competing smelting and hydrometallurgical processes is that any mercury entering the process from the sulphide concentrate feed is recovered rather than reporting directly to waste or polluting the environment. Outside the process circuit itself, the Intec Copper Process has several notable environmental advantages over competing technologies: Flexibility of scale with low capital and operating costs allow operation directly at the mill head at production rates as low as 15 000 tpa, significantly reducing environmental impacts from concentrates transport. The ability to handle low-grade and dirty concentrates often allows greater recoveries in the mill, improving both the economics and environmental acceptability of a mine by reducing metal losses to tailings. A comparison of the total energy requirements of the Intec Copper Process with those of competing technologies shows that the Intec Copper Process has the lowest energy consumption of any of the known hydrometallurgical processes, as well as improving upon pyrometallurgy under appropriate conditions.

Jan 1, 2002

By John A. Meech

In many multiple-sulphide ore flotation processes, it is desirable to concentrate each of the various valuable minerals into separate products. Although there are a variety of techniques and chemicals which can be employed to produce efficient separation, two basic types of flotation circuits are used: The Differential Flotation Circuit and the Bulk/Selective Flotation Circuit. Figure I depicts these two circuits in their most simplified form for the production of two concentrates and one tailing product. If three or more valuable minerals are present in an ore, a variety of combinations or extensions of these two basic flowsheets may be used. There are many examples of the successful application of both circuit types with ores containing similar minerals. Each ore must be tested individually to decide on which circuit is best. However, the following general statements can be made: 1) The differential circuit will have higher capital and operating costs than will the bulk/selective circuit due to higher rougher capacity requirement. 2) The differential circuit will achieve better separation than will the bulk/selective circuit since it is usually easier to reactivate a depressed mineral than it is to depress an already active mineral. If the deactivation problem with the bulk/selective circuit . . in can be overcome, there is a clear economic argument for its use

Jan 1, 1977

By W. H. MacLean, P. J. MacGeehan, A. J. Bonenfant

"The Norita deposit is an Archean volcanogenic-exhalative Fe-Zn- Cu sulphide accumulation of the distal type, consisting of six overlapping stratiform massive sulphide lenses, interbedded with exhalative cherts and thin pillow basalt flows, which formed down-topographic-slope from a mineralized hydrothermal discharge zone in the Norita rhyolite. The sulphide lenses are of two main types: channel-fill deposits such as the Main Zone and Main Zone West, composed of intricately layered pyrite-sphalerite-chert, which formed by the dynamic down-slope transport and re-deposition of sulphide exhaled from the vent zone; and stratiform lenses such as the Cu-rich A-Zone and A-Zone West, which formed even farther down-slope, in basinal depressions flanking the dome. Although rotated into a vertical position by folding, the sulphide lenses are otherwise essentially undeformed and contain a wide range of preserved primary sedimentary structures.Of prime importance in exploration has been the recognition that the formation of the distal sulphide lenses was controlled by submarine paleotopography, and that it is possible to approximate the sea-floor paleotopography using variations in the thickness and strike extent of the interbedded volcanic rocks. This has proved to be a major and successful exploration tool in the search for further sulphide mineralization in the mine, and offers prospects for use elsewhere in the district."

Jan 1, 1981

By Lawrence D. Hoy, Armand Savoie, Pierre Trudel, Ghislaln Tourigny, Robert Crépeau, Lao Kheang

"Recent fluid inclusion and stable isotopic studies in the Bousquet mining district have identified spatial and temporal trends in mineralization temperatures and isotopic compositions of vein forming fluids. These studies indicate that mineralizing temperatures were lower during sterile vein formation (T=200-300°C) than during auriferous vein formation (T = 300-400°C), and that 8 80 values of minerals in sterile veins are lower by 1-3 %o than in associated auriferous veins. In addition to these small differences in 8 80 within individual mineralized zones, a trend of increasing 8 80 values of vein minerals is identified from west to east across the district, with 8 80 of vein minerals increasing by4-6%o. Quantitative evaluation of potential causes of the observed oxygen isotopic shifts suggests that the best explanation for these trends is the derivation of the mineralizing fluids from an isotopically inhomogeneous terrain in which the fluid isotopic compositions are controlled by the isotopic compositions of the host lithologies due to high rock/water ratios in these source terrains. This inhomogeneous source terrain is interpreted as consisting of the Blake River Group volcanics (low 8 80 component) and the adjoining Kewagama and Cadillac Group sediments (high 8 80 component).The genesis of the hydrothermal fluids is attributed to crustal heating during metamorphism, with the fluids representing metamorphic dewatering products and/or fluids incorporated during deposition and burial of the host lithologies. Ductile deformation of the fluid source regions impeded fluid migration during metamorphism; the ultimate release of the mineralizing fluids is attributed to the penetration to depth of vertical structures (fault and shear zones) during brittle deformation of the crust under waning metamorphic conditions. Release of fluids caused by this puncturing of the closed hydrologic source regime produced a localized zone of lowered pressure, drawing fluids from adjacent lithologies toward the fault zones, feeding further hydrothermal fluid upwelling. Au mineralization resulted from chemical processes related to cooling of the hydrothermal fluids and/or interaction with lithologies encountered during fluid migration; the level of crust at which Au deposition occurred likely migrated upward with the thermal anomaly caused by hydrothermal fluid upwelling, eventually reaching the level of current exploitation. Cessation of fluid flow along the vertical structures is attributed to inadequate recharge of the hydrothermal systems as fluid contents were depleted in the lithologies proximal to the vertical structures and/or sealing of the structures by mineral deposition."

Jan 1, 1990

By Larry G. Stolarczyk

An underground medium frequency radio communications system has been developed and demonstrated in both coal and metal/non metal mines. The principal advantage of the system are its low cost, ease of installation and the ability to provide radio communications to any vehicle or roving miner in the underground mine. The system provides high quality radio coverage in work areas and along manways with existing "wireplant" conductors. The conductors include AC power cable, telephone cable, metal water and high pressure air pipes, etc. The measured communications range from the base station to mobile transceiver exceeded 12,500 ft without the use of repeaters. A cost/benefit analysis shows that the initial cost of the equipment can be returned 2-6 times in the first year. The mine productivity increases by approx. 13 percent.

By Peng-Fei He

CT scanning tests were used to detect the pre-existing fracture systems of 36 cubical coal blocks. Systematic procedures were developed to construct the pre-existing fracture geometry of the cubical coal blocks from the CT images. Based on the constructed fracture systems, the fracture tensor components were calculated to capture the directional effects of the fracture system of the cubical coal blocks. The effect of the pre-existing fracture system and different combinations of confining stresses on the strength and deformability of coal masses was investigated at a preliminary level using some true triaxial test results and the computed fracture tensor components. The strength of the cubical coal blocks in the loading direction was found to be closely related to the summation of the fracture tensor components in the two perpendicular directions to the loading direction and the level of the confining stress system. The deformability of cubic coal blocks in the loading direction was found to be correlated to the fracture tensor component in the same direction.

Jan 1, 2014

By Yu. A. Polozov, V. A. Lagunov, O. Yu. Lushinkova, Yu. I. Svirskiy, Eh. Ya. Kipko, Roy A. Williams

The integrated grouting method, as developed in the USSR, can be applied to multi-purpose operations that in¬clude the construction of shafts, drifts, and tunnels. In the USSR it has been used extensively for primary and final grouting of underground excavations in mining engineering and civil engineering projects. It has been applied in projects where the primary objective was waste contain¬ment. It has also been used to control subsidence beneath buildings located over mined out openings and to eliminate the flow of ground water beneath and around dams. This chapter considers several industrial applications in the form of case histories. The word integrated means that grouting activities were integrated with other activities. 9.1 CASE HISTORIES OF GROUTING FRACTURED ROCK WHEN SINKING VERTICAL SHAFTS The integrated grouting method has been applied most widely in the sinking of vertical shafts in fractured, satu¬rated rock. As explained previously herein, grouting oper¬ations preferably are carried out from the ground surface where they are integrated into the schedule of the setting up of the shaft excavation and construction equipment. Such integration reduces significantly the length of time required for preparation of shaft construction and the time required to sink the shaft. The increased efficiency is achieved by the elimination of cementing operations from the working face upon the penetration of each aquifer. Concomitantly, labor and energy consumption during shaft sinking are minimized because the more complicated and labor-intensive work is carried out at the surface section. Table 9.1 presents the essential parameters for the pri¬mary grouting of saturated fractured rock carried out by the integrated method when sinking vertical shafts. 9.1.1 NAGOL'CHANSK MINE NO. 1-2, VENTILATION SHAFT NO. 1 The integrated method of grouting saturated fractured rock through holes drilled from the surface was employed for the first time during the sinking of vertical shaft No. 1 at the Nagol'chansk mine No. 1-2 in the Don Basin. The shaft had an inside diameter of 6 m. The grouting operations commenced by using cement grout with a density of 1.7 to 1.8 g/cm3. Clay-based grout with cement and other addi¬tives was used only for grouting the aquifers where cement grout proved to be ineffective. By the time of the grouting of ventilation shaft No. 1, hydrodynamic analytical methods were well in hand and a preliminary method for designing isolation curtains had been developed. A method for designing and drilling ori¬ented directional drillholes whose natural curvature re¬flected the characteristics of the fracturing had been devel¬oped. The majority of the necessary equipment had been developed. Clay-based grouts were considered ready for the first industrial application. The hydrogeologic environment at the Nagol'chansk mine No. 1-2 contained many prolific aquifers. According to monitored drillhole data, the total expected inflow into ventilation shaft No. 1 without grouting was 425 m3/hr. This large ground water inflow rate was supported by in¬formation obtained during the sinking of the main shaft and from the auxiliary shaft of the "Nagol'chansk" mine No. 1-2. The construction contractor at the site began sinking ventilation shaft No. 1 to a depth of 217 m by cement grouting from the working face of the shaft. As a result of the cement grouting operations, the inflow of water was reduced from 105.4 to 40 m3/hr over this 217 m interval. However the residual inflow rate caused the shaft to become functional at the depth of 217 m. At this point all grouting operations were transferred to SPETSTAMPONAZHGEOLOGIA (STG) in an effort to improve the operation. In order to carry out preliminary grouting from the sur¬face, seven drillholes were installed and designed for re¬ceiving grout. Drilling was implemented by the ZIF-1200A rig to the design depth of the shaft. The geometry of the principal fracture system that would be encountered by the shaft was estimated while breaking ground using informa¬tion from other shafts. The drillholes were arranged so that they cut through the aquifers uniformly around the shaft. Some preference was given to intercepting the aquifers that were located along the rock dip above the shaft bottom because it was established that the ground water in the aqui¬fers flowed downward along the dip into the cone of de¬pression of the shaft. Investigations in the drillholes using the DAU-3M flowmeter described in Chapter 3 showed that all of the nine aquifers were penetrated by the time a depth of 690 m had been reached. The flowmetric information showed that aquifer numbers [1 and 2], [4, 5, and 6], and [7 and 8] had practically identical hydrogeologic characteristics; so it was decided to connect them into combined stopes and inject the grout using five stopes instead of nine. Table 9.2 shows the results of the grout injection into the aquifers. The volume of grout pumped into the separate intervals through single holes varied from 204 to 15 m3. The trend of grout consumption showed a reduction of grout consump

Jan 1, 1993

By Dong Yang, Jicai Han, Xiaoshuang Guo, Tichang Sun, Zhengyao Li

Straw charcoal was used as a novel reductant in the co-reduction roasting of laterite ore and red mud. The influence of the reductant dosages, co-reduction temperature, and co-reduction time were investigated. Results showed that a powdered ferronickel product of 1.81wt.% nickel and 81.40wt.% iron was obtained under straw charcoal dosage of 15wt.% and roasted at 1250 °C for 80 min. The corresponding recovery of nickel and total iron were 97.21wt.% and 98.87wt.%, respectively. The performance of straw charcoal was also compared to that by the conventional coke reductant. Results showed that ferronickel particles using straw charcoal were larger in size than coke, mainly due to the lower S (0.14wt.%) and ash content (3.80wt.%), which could not hinder the agglomeration of the ferronickel grains, and lower melting point that producing liquid phase and promoting ferronickel particles growth. In addition, the gasification reactivity of straw charcoal is better than that of coke, which is beneficial to the iron oxide reduction and the recovery of iron. To conclude, straw charcoal is a promising reductant for the co-reduction of laterite ore and red mud in coal-deficient areas.

Aug 12, 2021

By W. C. Risley

IN the year 1933, workings at the No. 12 colliery, New Waterford, of the Dominion Coal Company, Limited, had advanced to a point where additional ventilation would soon be required. Furthermore, development of the mine was being .made to practically double the output of coal. Consequently, authorization was given to sink a shaft to the workings. The location selected is on the west side of the mine, about one and a half miles from the slope mouth and on the Low Point shore, some 250 feet from high-water mark. The cliff at this point is 40 feet high, with the land rising on a slope of about 5 per cent to the shaft. Figure 1 shows the location with reference to the colliery workings. The strata through which the shaft was to be sunk were pretty well known from diamond-drill borings previously made, and it was believed that water would be encountered to a depth of at least 250 feet to 300 feet. It was planned, therefore, to seal-off the water by the pumping of cement grout under high pressure. This method was considered to be the most economical means of handling the water during sinking, and would also result in a permanently dry shaft. Figure 2 shows a cross-section of the strata. The permanent shaft hoist and fan were to be electrically driven, which would necessitate the construction of a transmission line between the colliery and the shaft; therefore, it was arranged that the sinking equipment would have electric drive. As all equipment, materials used in construction, and supplies were to be hauled by truck, a roadway 1,800 feet long was required from the main highway to the site, and as heavy loads would be hauled during the fall, winter, and succeeding spring, this road was ballasted with 18 inches of clinkered material obtained from a burned culm bank one and a half miles distant. Actual field work was begun by the Company's own forces early in October. Work preparatory to sinking-in addition to the roadway and trans-mission line-included the construction of the reinforced-concrete shaft collar and fan drift connection; the permanent timber headframe (which was to be used also as the sinking headframe); the permanent reinforced-concrete building to house the shaft hoist and fan drive; and the erection of temporary buildings and installation of all equipment for sinking and lining the shaft and for cementation.

Jan 1, 1936

By Wadaw Cybulski

The problem of coal dust explosions: Basic information about the explicability of coal dust: Main parameters of the explosibility of coal dust; Main parameters of the explosibility of coal dust; Limit explosibility of coal dust and its dependence upon the basic parameters: The effect of the initiator of the explosion-on-the explosibility of coal dust: The dependence of the course of an explosion upon the distribution of coal dust in the workings, upon the dimensions of the workings, the length of zones, the direction of the development of explosions and upon the content of firedamp in the air of the workings; Con-sequences of coal dust explosions; The ways of suppressing the danger of coal dust explosions; Stopping the developed coal dust explosions; The explosibility and explosions of lignite dust of xylite structure

Jan 1, 1975