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By Z. Kowalski

In the defunct Gorka quarry, there are situated both a waste disposal site of an area of 6.7 ha and containing approximately 1 million ton of high aluminium wastes and a pond of an area of roughly 3 ha, up to 15 m deep, with about 400000 m3 of polluted water. The reservoir is fed with the filtrates arriving from the disposal site mentioned, at a rate ~130 m3/day. An alarming phenomenon is the subsidence of the pond bottom and the infiltration of solutions into the Triassic and Jurassic water resources, assessed at ~40 m3/ day. The basic problem related to the effluents in the Gorka pond is their high alkalinity (pH 12-13.4) and variable pollutants whose level increases appreciably with the pond?s depth. The solution offered provides pumping out and treatment of about 500000m3 of effluents retained in the Gorka reservoir in an about yearly time. Effluents would be treated in a reverse osmosis plant, according to a process so far verified on a quarter-commercial scale. The treatment process by-product would be discharged into the Ropa stream. The brine solution (containing about 25% NaCl), whose amount may reach up to100000m3, might be supplied to the Debiensko Mine which produces technological salt from salt-bearing mine waters. Similarly, there would be provided an intake of underflows of infiltrates (by means of deep wells, and possibly girdling ditches) arriving from the high aluminum site, and such effluents would be subject to ongoing dilution in water courses. It would be possible to test the system of underflows reaching the Gorka reservoir by means of geochemical identification, previously performed. The next stage following the pumping, would be the utilization of approx. 50 000m3 of bottom slurry. Highly alkaline slurries would be utilized in the production of self-solidifying mixtures. Slurries would be pumped out and temporarily dumped on drying beds, located nearby the Gorka reservoir, and later on used successively for making self-solidifying mixtures in a plant similar to a concrete making seat. Those mixtures would be applied for sealing the bottom of the Gorka reservoir and the necessary part of the edges of the defunct quarry. An alternative solution is the utilization of slurries in the facilities of the Trzebionka Mining Works and the application of self-solidifying mixtures, made in the Trzebionka Works, for sealing the reservoir bottom.

Jan 1, 2006

1. Book reviews Underwater tools, by Donald J. Hackman and Don W. Candy. Columbus (D.S.A.), Battelle Press, 1981. 152 pp. $32,95. Reviewer: B. Schmitz Uranium exploration case histories. Proceedings of Advisory Group Meeting Vienna, IAEA, 1981. Price: Austrian schillings 630. Reviewer: E. C. Kostlin Jet cutting technology, a review and bibliography, edited by Robin Brown. Published by BHRA Fluid Engineering, Cranfield, Bedford MK430AJ, England. Reviewer: D. O'Beirne 2. New Books Proceedings, Twelfth Congress of the Council of Mining and Metallurgical Institutions, by H. W. Glen (editor.) Johannesburg, The South African Institute of Mining and Metallurgy, 1982. 2 vols., 1060 pp. R78,85 (air-mail RI53,85). Amborite. Johannesburg, De Beers Industrial Diamond Division (P.O. Box 916), 1982. 6 pp. Free of charge. 3. New Journal BHRA Fluid Engineering has announced its new journal World Ports and Harbour Abstracts and the associated newsletter World Ports and Harbour News, which are issued from January 1983. The Abstracts is published bi-monthly and costs £60 in the V.K. and E.E.C., and £72 elsewhere. 4. Mintek reports Report 2134D On-line analysis, by X-ray diffraction, of fluorsparcontaining slurries. (First published 30th Sep., 1981.) Report 2135D A preliminary on-line analysis, by X-ray diffraction, of pyrite-containing slurries. (First published 30th Sep., 1981.) Report M28 The measurement of flow velocity distribution in fluids. Report M34 The determination, by atomic-absorption spectrophotometryusing electrothermal atomization, of platinum,palladium, rhodium, ruthenium, and iridium. Report M45 Two computer programmes for the calculation of massand heat balances in the production of high-carbon ferromanganese. Report M50 The development of an on-line gold analyser. Report M54 Major amounts of mercury in native gold from upper Witwatersrand sediments. Report M55 A new robust method for the treatment of analytical data. Report M57 The hydrometallurgical treatment of liquors generated by the leaching of a low-grade matte. Awards for inventions For inventing devices of exceptional value to the mining industry, two mining men were presented with awards in Johannesburg on 25th November, 1982, under the Chamber of Mines' Ideas and Suggestions Scheme. Mr W. W. Malan, President of the Chamber of Mines, made the presentations.

Jan 1, 1983

By Derek H. C. Wilton, David T. W. Evans

"Abstract - The Midas Pond gold prospect, central Newfoundland, is hosted by sheared and altered felsic and mafic pyroclastic rocks of the Tulks Hill volcanics, Victoria Lake Group, and has been trenched and tested by 19 diamond-drill holes. Gold values are sporadic, with the best diamond-drill intersection assaying 7.3 g/t over 0.9 m, and the highest grade channel sample containing 14.7 g/t over 1.15 m. Alteration and mineralization are confined to a 200 m wide brittle-ductile shear zone. This shear zone formed in response to regional D1 deformation, and the shear zone fabric parallels the regional S1 foliation. D2 deformation resulted in broad Z-shaped flexuring of the shear zone. The deformation was probably related to the ca. 396 Ma to 420 Ma Salinic Orogeny.Advanced argillic alteration and an extensive halo of Fe-carbonate and pyrite surround the gold mineralization. Alteration minerals include pyrophyllite, paragonite, quartz, plagioclase, chlorite, fluorite, Fe-carbonate, and pyrite. As such, the alteration seems to reflect combined epithermal (argillic) and mesothermal (carbonate) mineralizing environments. The argillic alteration and fluorite represent a regional hydrothermal effect on the felsic volcanic rocks; whereas, the Fe-carbonate and pyrite represent a mesothermal alteration associated with auriferous quartz veins. The Au, which is spatially associated with pyrite, occurs in three structurally-controlled vein sets. These veins are confined to the contact between a highly deformed breccia (the banded mafic tuff) and structurally overlying, altered felsic tuffaceous rocks. The vein sets include: V1 boudinage veins that parallel the shear-zone fabric (C-shear veins), and V2 and V3 extensional fracture veins which are controlled by an S2 fracture cleavage. V1 veins are the earliest and contain the lowest concentrations of Au. V2 and V3 veins are concentrated within the hinges of the D2 flexures.Pb isotope data for galena separates indicate that the Pb in the vein systems was probably of local derivation. Sulfur, C, and O isotope data for mineral separates are similar to data for typical mesothermal lode gold deposits. Fluid inclusion data suggest the presence of at least three hydrothermal fluid types in the region, viz., (1) an early saline fluid which produced the argillic alteration, (2) a CO2-bearing fluid directly related to the auriferous mineralization, and (3) H2O-NaCl fluids which flowed through the Midas Pond shear zone at a lateral distance from the zone of mineralization. The mineralizing fluids are interpreted to have been CO3 -2-rich types which originated through metamorphic dehydration and decarbonation of mixed island-arc and continental crustal rocks. Gold precipitation resulted from the reaction of these fluids with the Fe-rich breccia unit. Pre-existing and simultaneously crystallizing pyrite served as loci for Au precipitation."

Jan 1, 2000

By D. C. Torre

The need for a successful, efficient and reliable continuous face- haulage .system for use with continuous mining machines has been apparent since the first continuous miners were introduced some twenty-five years ago. Some thought and effort were given to the development of continuous haulage over the last twenty years or so. Some initial success was experienced in this area by The Long Company with a system involving bridge conveyors, as far back as the late 1950's. However, extensive success with continuous face haulage for use behind modern, high product¬ion, continuous miners escaped the industry until the early 1970's. It is true that continuous face haulage was successfully introduced into low seam mining, generally in seams ranging under 1.07 m (42") in height, in the 1960's, for use with relatively low production, low seam, auger type continuous miners. However, those available systems which have achieved success in the lower seams, where relatively low capacities are adequate, are not directly applicable for use with today's medium and high seam miners with their standard production capacity of 9.1 - 10.9 mtpm (10 - 12 tpm). Since the late 1960's and early 1970's, several machinery manufacturers, and at least two major coal companies, have devoted considerable expense, time and effort to developing truly successful continuous face haulage for use in coal seams ranging upward from 1.2 m (4') in height. Continuous face haulage equipment has been developed in the last five years employing a variety of transportation media and design concepts, including some which have received wide publicity such as those employing hydraulics, and pleated belt. The truth of the matter is, however, that the only universally applicable and widely successful continuous face haulage equipment introduced to date has been that employing a combination of bridge conveyors and mobile, self-propelled, bridge carriers, connected at the inby end to the tail of the high production continuous miner, and at the outby end attached directly to the panel conveyor (Fig. 1 & Fig. 2) or, on occasion, to a short, self-contained conveyor which, in turn, empties directly onto the panel conveyor. While these bridge type continuous haulage systems are available with both belt and chain conveyor components, the most widely applied and successful system in seams above 1.07 m (42") in height is that offered by The Long-Airdox Company. The standard unit for use with 9.1 - 10.9 mtpm (10 - 12 tpm) continuous miners is normally equipped with belt. However, they have built, and successfully applied, versions of this system employing chain conveyors, (Fig. 3), mainly for operators having extremely difficult mining conditions, including excessive amounts of water

Jan 1, 1977

By H. S. Chiang, Y. Luo, S. S. Peng

"IntroductionThe way in which the particle sizes distribute in the airborne coal dust plays an important role in the transportation of the dust in the mine air and human respiratory systems. Though the particles of a size less than 10 microns could be inhaled, the real distractive portion of the dust is that the sizes range from 1 to 5 microns since only they can reach the lungs and be retained there. Those with an equivalent diameter larger than 5 microns will settle on the membranes of the branchial tubes(1) while those with diameter less than 1 micron can either be exhaled out of the lungs or the settled quantity is so insignificant that does very little harm to the lungs. In the transportation process, the finer dust will fly farther away from the sources than the coarser ones. Thus once the fine dust is generated and becomes airborne, it will spread out to a larger area.In this paper, the coal samples collected from two longwall faces were analyzed statistically for size distribu¬tion. A mathematical model is proposed and the effects of different conditions on the dust size distribution were analyzed using the mathematical model.Sampling and Statistical AnalysesSampling Instrument and ProcedureThe instrument used in the sampling were Mine Safety Appliance (MSA) personal sampling units equipped with 10-mm nylon cyclones, pre-weighed MSA PVC filters and constant-flow pumps.Two longwall operating faces were selected for the dust sampling. The two faces were located in two coal mines ex-tracting Pittsburgh seam. The double-ended ranging-drum shearer applying the unidirectional cutting sequence was employed. The two-leg shield supports were used for face support. The width of the two faces was about 215 meters.In each of the two faces, four sampling stations were set up. The stations were located at the cross-cut immediately outby the headentry side, the 15th support (25 m), 65th support (107 m) and 111th support (182 m) from headentry T-junction, and numbered as #1, #2, #3, and #4 station, respectively (Figure 1). At each sampling station, a group of three sampling units was hung at proper position. At cross¬cut (#1) station, the units were hung about one foot under the roofline in the center of the cut (Figure 2a), and at the stations in the face, the units were position near the hydraulic legs and one foot below the support canopy (Figure 2b).One complete set of 12 samples was collected for each day. The sampling were performed for five consecutive days. All the samplings were conducted in the morning shifts. The time duration of the sampling for each day was about five hours. During that time, the shearer might complete 5 to 7 mining cycles."

Mar 1, 1989

By K. J. Henderson

INTRODUCTION Shrinkage mining at Inco Ltd.'s Crean Hill mine can best be defined as a horizontal breasting method utilizing the broken ore for the mining floor and to provide wall support, with draw from a series of boxholes or draw¬points at the base of the stoping block. In choosing to shrink a given block of ore, the initial investigation centers on the ore configuration and grade, the expected ground conditions in the stoping area, the expected profit compared to alternative mining methods, the future production requirements, development lead time, the availability and type of equipment in the area, and the availability of experienced miners. These factors are weighed against other mining methods and shrinkage mining is chosen when the advantages outweigh the dis¬advantages. These will be discussed in more detail later. At Inco Ltd.'s Crean Hill mine, shrinkage mining makes up 17% of total mine production and is decreasing due to the advent of advanced blasthole techniques, notably in-the-hole drilling. Shrinkage is an adaptable mining method that is currently employed in narrow irregular ore zones, as undercuts for in-the-hole blast¬hole stopes, in ore bodies that do not extend from level to level, and in areas where quick production is required with a minimum of capital outlay or minimum develop¬ment. Two major types are currently in use which are classified on the basis of type of draw system employed. These are the utilization of load-haul-dump (LHD) equipment drawing from drawpoints, and the slusher and train method drawing from boxholes into a slusher trench and loading into a train. A third method being phased out is drawing from chutes directly into a train. In all three methods the actual stoping operation remains the same. In the initial planning of a mining block, the ore body is first located using an exploration diamond drill¬ing program. This is essentially done at a 130-m (400¬ft) spacing and it gives an indication of the amount of ore, grade of ore, the configuration of the ore body, as well as other geological features such as ground condi¬tions which will affect future mining. This drilling is done from an exploration drift 165 m (500 ft) long in the hanging wall and covers a vertical extent of 330 to 670 m (1000 to 2000 ft). However, these data are not accurate enough for detailed planning. When the de¬cision has been made to mine, a preliminary develop¬ment drill drift is driven in such a position that it can serve the dual purposes of allowing the ore body to be drilled off in detail, or of later serving as a haulage or access drift. Diamond drilling from this drift is usually on a 15-m (50-ft) grid spacing and gives adequate information for detailed planning. Utilizing the drill data, a decision is made on the mining type. For a shrinkage operation, the stoping block is ideally 49 m (150 ft) in length and a maximum of 13 m (40 ft) in width. Lengths of 65 m (200 ft) are considered a practical maximum because of the length of time the back of the stope remains open during the min¬ing of a cut. A minimum length of 13 m (40 ft) is used because the amount of time spent in opening a new cut increases in comparison with the total stoping time. The maximum width of 13 m (40 ft) is used because ex¬perience with the ground indicates that generally stope backs tend to be unstable over greater widths. However, widths of 33 m (100 ft) have been mined. A minimum width of 5 m (15 ft) is used to allow an effective draw, but in unusual cases, a width of 3 m (10 ft) has been mined successfully. The dip of the ore should not be less than 1.04 rad (60°). Angles flatter than this will not pull properly resulting in inefficient mining because of ore hang-ups on the footwall side of the stope and the need for large stagings on the hanging wall. In considering shrinkage, mining ground conditions play an important role. Strong shearing, cross faulting, or dikes are avoided. These usually cause dangerous mining conditions and an alternative safer mining method which will give adequate support to these adverse ground conditions is employed. Production requirements are taken into considera¬tion when planning a shrinkage mining program. In it¬self, shrinkage mining is capable of producing ore in a relatively short time with little capital development. If quick production is required, this method can produce ore rapidly as all that is required is that the draw system be completed and the miners with jacklegs and stopers produce a limited amount of ore. The obvious disadvan¬tage is that not all the ore broken can be pulled from the stope immediately. However, a mining horizon with sev¬eral shrinkage stopes in different stages of production can adequately maintain scheduled tonnage require¬ments, the cycle being stopes under development, stopes actively mining, and stopes completed and pulling. When the decision has been made to use shrinkage mining in a given block, the haulage and draw system is driven. The proximity of the orepass system dictates whether the system is to be a LHD operation or a slusher trench operation. If the orepass system is in close prox¬imity to the stoping block, LHD equipment is employed because of its ability to operate efficiently over short haulage distances. A long tram to the orepass usually dictates a slusher trench operation with haulage by train. Internal orepass systems are used effectively with LHD equipment with a load-out and retramming taking place on a lower horizon. In addition, stoping blocks in older areas of the mine are developed track because of the capital cost involved in converting an existing level to trackless mining. DRIFT DEVELOPMENT USING LHD EQUIPMENT The haulage drift for a draw system using LHD equipment is driven parallel to the ore body in the foot¬wall at a sufficient distance to allow an 11.5-m (35-ft) drawpoint to be driven into the ore (Fig. 1). The actual position of the drift depends entirely upon the ore con¬figuration and the drift size depends on the equipment to be used. Drift sizes of 4.6 x 4.1 m (14 x 121/2 ft) for an St-4 and 3.8 x 3.6 m (111/2 x 11 ft) for an St-2 meet company and government clearance standards and allow sufficient space to accommodate a ventilation pipe. The

Jan 1, 1982

By John Babcock, Bryan C. Smith

In a partnership between Luck Stone Corporation and Loudoun Water, a 42” (1.067m) steel pipeline was constructed along the perimeter of an extensive permitted reserve of diabase rock Luck Stone was about to begin mining. Loudoun Water provides drinking water and wastewater services to over 75,000 households in Loudoun County, Virginia. In discussions to establish the 30-foot (9.1 meter) easement for Loudoun Water within the permitted reserve of diabase rock, Luck Stone and Loudoun Water agreed on the following basic conditions: • Luck Stone could execute production blasting as close to the easement as physical conditions of overburden depth permits. We estimated the minimum distance to be about 75-feet (22.86 meters) from the centerline of the pipe. • Loudoun Water would install strain gauges on the pipe every 500-feet (152.4 m) along the entire length of the future reserve area. This resulted in the installation of 18 strain gauges. • Luck Stone would contract a third-party vibration monitoring company, agreed to by Loudoun Water, to monitor the strain gauges and provide seismograph reading at appropriate locations along the pipe as mining progresses. All information would be provided to Loudoun Water. • Luck Stone’s blasting will be conducted in a manner that ensures the recommended strains are not exceeded.

Jan 1, 2019

By P. W. Jensen

Gold deposits in Interior Alaska are commonly coincident in time and space with reduced, mid-Cretaceous granitoids. The True North deposit of the Fairbanks district is several km from such granitoids, but yields mineralization ages indistinguishable from the Fort Knox granite-hosted deposit. True North, however, resembles sediment-hosted epithermal Au, demonstrating the variability of intrusion-related Au deposits in metamorphic rocks. The True North deposit (14M tons at 1.5 gpt) is hosted within a carbonaceous thrust complex separating eclogite facies from amphibolite facies rocks, located > 5 km from mineralized granite outcrops. The ores contain anomalous Sb, As, and Hg (stibnite is the most common sulfide) with ankerite-sericite alteration in steep NE-striking ?feeder? shears and low angle thrust shears. Nearest hornfelsed rocks are located > 1 km away, and across several high-angle faults. Fluid inclusion data indicates trapping temperatures of ~250C from low-density, low-salinity fluids at P < 0.5 kb. Au was originally present as solid solution in sulfides; minable ores are restricted to the oxidized upper 100 m of the deposit. The considerable differences between this deposit and the temporally equivalent Fort Knox intrusion-hosted deposit indicate no one description adequately characterizes this deposit type.

May 1, 2003

The re-use of waste water from coal preparation plant tailings is always desirable, but is generally problematic in terms of solid-liquid separation. Dewatering and disposal of coal slurry at a high percentage of moisture poses a major technological and environmental challenge to mining and mineral processing industries world-wide. Characterization and flocculation of these coal tailings reveals that the separation of fine coal tailings depends on factors such as size distribution, slurry viscosity, zeta potential, pulp density, circuit design, plant operating parameters, and the behaviour of solid particles in an aqueous medium. The present study was carried out on -600 mm coal tailings collected from a coking coal preparation plant of Jharia Coalfield, India, using anionic flocculent. Characterization studies revealed that the zeta potential was highest (-44 mV) at a pH of 9.0, the size range of the particles was 1.1?700 m, and the average ash content was 62.57%. Flocculation studies were performed at different solid concentrations, pH values, and polymer dosages. The results indicated an efficient settling rate of 178.15 mm/min at 8% solid (w/w) at a flocculent dosage of 32.50 g/t solid. The turbidity of the clear liquid obtained was 7.42 NTU.

Jan 1, 2014

By R. B. Rao, B. Das, J. K. Mohanty, D. S. Rao

Physical, mineralogical and geochemical characterization, along with some beneficiation approach to pre-concentrate the Ni content of the chromite overburden material of the Sukinda Valley, India, have been carried out in detail. Mineralogically, the overburden consists of geothite, quartz, hematite, magnetite and secondary silicates in varying proportions. A specific Ni mineral phase is absent in this material with the Ni and Fe being mostly enriched in finer fractions. Distribution of Ni and Fe in various size fractions and increase of Ni with increase in Fe (goethite content) indicate that goethite is the major Ni-bearing mineral phase. Quantitative estimation of goethite from chemical analysis and weight loss data indicates that the goethite quantity varies from 40% to 90%. Electron microprobe analysis (EPMA) indicates that Nirich and Ni-poor goethite are present in the overburden material. A sequential dissolution study indicates that more than 50% of the Ni are present within the goethite lattice as a lattice substitution. The results of pre-concentration studies (size classification, gravity, flotation and magnetic separation) indicate that Ni is richer in finer fractions. Further, the enrichment of Ni content is limited to 1.1% with 50% recovery by classification followed by hydrocyclone separation of the below 53 m m fraction.

Jan 1, 2000

By A. René Silva

The separation of Fe(III) from simulated Ni(II) solutions using different ion exchange resins was studied. The resins tested included cation exchangers with sulfonic and carboxylic groups and chelating resins with iminodiacetic, aminophosphonic, and mixed sulfonic and phosphonic groups. The dosage of resins was varied from 0.1 to 0.5 g/mL to study the impact of resin type and dosage on iron separation from simulated nickel leach solution containing 25 g/L Ni(II) and 25 g/L Fe(III) at pH 1.5. Also, the effects of different metal ion concentrations, Fe (III):Ni(II) ratios in solution, pH, and adsorption times were investigated. Experimental results demonstrate that all the chelating resins selected favored the adsorption of Fe(III) over Ni(II) with high Fe(III) loading. Soft cation exchange resins also favored the adsorption of Fe(III) over Ni(II), but no significant Fe(III) loading was observed. In contrast, strong cation exchange resins did not show adsorption preferences. The most efficient separation was observed using resin with aminophosphonic groups at resin dosage of 0.2 g/mL, having a Fe(III) adsorption over 94% and Ni(II) adsorption below 5%.

By K. Hanna

Early identification of adverse geologic conditions and high stress concentrations during the high-speed extraction of coal using modern mechanized long wall systems is a major challenge facing the mining industry. Often, unexpected geologic anomalies that affect rock mass quality are encountered more rapidly, leaving the mine operator with insufficient time or information to deal with the changing ground conditions. The RockVision3DTM seismic tomographic system, developed by NSA Engineering, Inc., offers substantial benefits over conventional geologic/stress-detection methods and has been used successfully in coal mines worldwide. Tomographic images, produced economically and in near-real time, are easy to interpret, allowing mine operators to deal with potential hazards on a day-to-day basis. This paper focuses on the application of RockVision3DTM technology for delineating hazards caused by stress and geologic anomalies. Case studies for two bump-prone coal mines will be discussed. For example, (1) in a German deep (>1,130 m) mine, the drilling-yield method was used to verify the effectiveness of this technology in detecting stress anomalies ahead of longwalling; (2) in a Polish mine (-500 m deep), two longwall panels 4 km apart were monitored to assess structural and stress-related factors contributing to differences in mining conditions and magnitude of structural failures.

Jan 1, 1999

By Zhong g. Shui-ping

A new type Pb-Ag (0.8wt%)-Bi (0~5.0wt %) alloy used as anode in electrowinning of zinc from sulphuric acid electrolytes was prepared. The electrochemical properties of this ternary lead alloys have been tested and compared with those of Pb-Ag (0.8wt %) anodes used in industry. Results show the anodic overpotential decreases with the increasing of Bi content in the alloys. When the content of Bi was 1.0 wt% and 5.0 wt%, the anodic overpotential is about 40?50 mV and 60?80 mV lower than that of Pb-Ag (0.8wt %) alloy, respectively. The formation of oxide layers on the surface of these alloys was traced by LSV methods and the composition of the anodic oxide layers was characterized by XRD and SEM analysis. The surface layer on the two type anodes examined is mainly composed of a?PbO2 and ß?PbO2. But different structure of the surface layer was observed, Pb-Ag (0.8wt%)-Bi (1.0wt %) has a more fine-grained structure as compared to Pb-Ag (0.8wt %). But the film on Pb-Ag (0.8wt%)-Bi (3.0wt %) alloy is rather friable, cracked and not tightly attached to the anode.

Jan 1, 2008

By Bastiawarman Rifki, Nugraha Nico, Priatna Anton

The Deep Mill Level Zone (DMLZ) is one of the underground mines operated by PT Freeport Indonesia (PTFI). DMLZ is located approximately 1,500 m below ground level, where dominant rock types are skarn and diorite. The DMLZ mine and layout was designed based on the previous mine located above ground, the Deep Ore Zone (DOZ). The DOZ mine, located approximately 1,000 m below ground level, was a proven success, with a maximum capacity of 80K tonnes per day. During the production ramp up, which started in 2015, the DMLZ experienced several seismic events. The most significant event occurred in January 2018 with a recorded magnitude of more than 2.0 Mw, which resulted in a 14-month delay of the undercutting until March 2019. The significant event in January 2018 required the DMLZ to apply several adjustments related to the mining plan to help ensure it could continue development and achieve production target safely. This paper will describe the changes in the DMLZ mine plan as a response to the seismic activity at the mine, the reason behind the decisions being made to change the mine plan and strategy, as well as strengths and weaknesses between the previous plan compared with the new plan.

Apr 26, 2022

By S. D. Maksimovic

The Bureau of Mines identified and evaluated major sources of quartz dust during overburden drilling at a surface coal mine, The relative contribution of each source to the total dust hazard was estimated, using personal gravimetric and instantaneous dust-sampling devices. No attempt was made to determine actual dust generation rates. The major sources of dust at the drill rig during dry drilling opera-tions were the collector dump, drill shroud leakage, and drill stem seal leakage, These sources contributed 90 pct of the respirable dust and 89 pet of the respirable quartz dust generated during drilling, The collector dump cycle (21.6 pct of total drilling time) accounted for 38 pct of the respirable dust and 41 pct of the respirable quartz dust, with peak dust concentrations reaching 68 mg/m3. Shroud and drill stem leakage contributed 28 and 24 pct of the respirable dust and 32 and 16 pct of the respirable quartz dust, respectively. A dragline and second drill rig contributed the remaining 10 pct of dust and 11 pct of quartz dust. Gravimetric dust samples collected inside the drill cab and outside, near the cab door, averaged 1.65 and 1.43 mg/m3, respectively, with 6 pct respirable quartz.

Jan 1, 1985

By Ann M. Hagni

The mineralogical, chemical, microtextural, and particle size character of electric arc furnace (EAF) dusts are being studied by reflected light microscopy, scanning electron microscope-energy dispersive spectroscopic analysis (SEM-EDS) and electron probe microanalysis (EPMA). The dusts consist predominantly of metallic iron, iron oxide, and calcium-iron-manganese silicate phases, together with smaller amounts of iron-manganese-zinc-calcium, iron-calcium-zinc-manganese, and calcium-iron-manganese silicates. The samples vary in their zinc contents. Chlorine occurs in phases containing iron, lead, zinc, manganese, and calcium. The dust particles range greatly in size from 1 to more than 300 ~m, but most particles are about 1 to 6 ~m. Larger particles. about 20-50 um across. also are present. They are predominantly spherical in shape, but some are cenospheres, and others are more angular particles, some of which are broken spheres and cenospheres. Their internal microtextures exhibit evidence of liquid immiscibility, exsolution, skeletal crystal growth, and oxidation react i on rims. Because EAF dusts have been classified by EPA as hazardous wastes, there is heightened interest in the treatment of these dusts to render them non-hazardous. The mineralogy of products of pyrometallurgical treatment of EAF dusts by flash furnace and pyrohydrolysis are currently under study.

Jan 1, 1991

By John C. Edwards

Carbon monoxide (CO) sensors are used in mine fire early-warning systems. Experiments were conducted by the U.S. Bureau of Mines with six diffusion-type CO sensors to determine their response to a reference CO concentration under both static and dynamic flow conditions. Static experiments were conducted with 0.5 L/min flow of 16 ppm CO in N2 through 8-and 12-L chambers. Dynamic flow experiments were conducted in two environmental chambers: One had directional flow capability of 1 m/s, and the second provided moderately nondirectional flow as high as 1.8 m/s. Reference CO concentrations of 13.5 to 19 ppm in air were used for the dynamic experiments. The results of the static experiments demonstrated that the sensors consistently underestimated the reference gas concentration. The dynamic experiment results showed the sensors provided reasonable agreement with the reference gas concentration when directional flow effects were not considered. Directional flow effects were evaluated for two sensors. One of the sensors showed a strong dependence upon the relative orientation of the diffusion tube with respect to flow direction. The second sensor did not show a dependence upon flow direction. The sensors&apos; directional flow effect is significant for their proper in-mine utilization.

Jan 1, 2010

By John H. Schissler

INTRODUCTION AND SUMMARY Complex copper-lead-zinc ore is mined from underground workings high in the Western Cordillera of the Andes mountains about 100 km in a stralght line north of the city of Arequipa in southern Peru. The Madrigal mine is situated in a steep-walled canyon into which the six operating levels "daylight". Production is from cut-and-fill stopes currently producing approximately 740 dry metric tons per day. The Santa Rosa vein, the most important producer in the district, outcrops for a distance of about 4 km, bearing N 75º W and dipping about 70º NE. In places the vein splits into two or three branches; widths of each vary between 1 and 4 m and total mineralized widths of up to over 8 m occur. Principal ore minerals are chalcopyrite, galena and sphalerite. Ore reserve grades for the Madrigal mine average about 1.45% Cu, 2.6% Pb and 5.0% Zn, 2.9 ozAg per metric ton. [Compania] Minera del Madrigal is a California limited partnership; the general partner and operator is [Compania] Madrigal, a wholly-owned subsidiary of Homestake Mining Company. HISTORY Shallow workings found on the Santa Rosa vein were probably made by the Spaniards during colonial times in their search for

Jan 1, 1977

By B. Amero, W. Petruk, D. R. Owens, D. R. Morris

Samples taken at different heights in the lead blast furnace of Brunswick Mining and Smelting corporation were studied to determine the phases present in order to interpret the reactions In the furnace. The main phases in the blast furnace feed are Pb3Ca2S13011, Zn ferrite (ZnFe204), and hardystonite (Ca2ZnSi207); rninor ones include lead and a Pb silicate. The temperature in the furnace varies from about 300°C at 4.2 m above the tuyeres to 1150°C at 1.1 m. There is no evidence of reactions in samples from points cooler than 800°C. Some reaction occurs between 950° and llOO°C and the lead sinter is nearly decomposed at about 1150°C. The first evidence of reaction was the decomposition of Pb silicate to lead and S102. At higher temperatures the Pb Ca2Si3011 decomposed to form lead and hardystonite (Ca2ZnS120rf. The Zn required for hardystonite was derived from the Zn ferrite which apparently re-equilibrated to a new Zn ferrite with less Zn. The decomposed sinter consists essentially of lead, hardystonite, wustite and a Zn ferrite with a low Zn content.

Jan 1, 1984

By H. H. Einstein

Swelling ground is one of the most feared problems in tunneling. Although usually not causing sudden failures during construction, the very large and longlasting deformations give contractors and designers the impression of facing an inevitable phenomenon against which there is no technical defense. Examples are concrete liners up to 1 m (3 ft) thickness that have been cracked or crushed, tunnel inverts that are heaving at 0.1 m (0.3 ft) per week during construction and continue to heave during decades. This paper summarizes essential components of design construction procedures where tunnels in swelling rock can be built to perform satisfactorily. It is important to point out that design and construction cannot and should not be separated. The way the tunnel will be built has to be an integral part of the design. Swelling can be defined as time dependent volume increase and is distinguished from squeezing, another time dependent phenomenon occurring in tunneling. Squeezing involves exceeding the shearing resistance while swelling usually does not. Swelling of rock masses can occur as swelling of the intact rock, swelling of wall rock, of gouge and of filler in the discontinuities of a rock mass, and as a combination of the above. Swelling mechanisms have been discussed elsewhere (ISRM 1983).

Jan 1, 1984