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By A. D. Davis

This project took a watershed approach in the evaluation of water-quality impacts from past mining activities and current restoration projects in the Bear Butte Creek basin, a small and well defined drainage in a heavily mineralized area of the northern Black Hills. The basin contains numerous inactive and abandoned mines as well as a currently active heap-leach surface gold mine. Bear Butte Creek normally loses all of its discharge to sinkholes in the Madison Limestone aquifer and thus can affect the water quality of this regionally important water resource. Water-quality data were analyzed, stream surveys were conducted, and two abandoned mine restoration projects were monitored to evaluate the effectiveness of the watershed approach and to identify practical remediation techniques that could be applied to other watersheds. A geographic information system (GIS) was developed to organize information on stream quality, abandoned mine locations, and other geologic and hydrologic data. The watershed approach toward assessing and reclaiming abandoned mines appears to be effective when compared to a point-source approach that might improve water quality at specific sites but not result in significant overall improvement to the watershed because of adjacent sources.

Jan 1, 1997

By Jeff LeBlanc, George Gutierrez, Carl Pitzer

The Bay Park Conveyance Project was one of the largest microtunneling and rehab projects in the United States at the time of completion. The completed project conveys treated water from the South Shore Water Reclamation Facility located in Nassau County, New York, which currently discharges an average of 50 million gallons per day (MGD) of treated water into Reynolds Channel, to the Cedar Creek Water Pollution Control Plant (WPCP) ocean outfall pipe. The pipe carrying treated wastewater was installed approximately 20–60 feet below the surface. The treated wastewater is conveyed via microtunnel construction of a 2-mile long force main from the South Shore Water Reclamation Facility to an existing aqueduct under the Sunrise Highway, the rehabilitation of an 7.3-mile stretch of the aqueduct via sliplining, and construction via microtunnel of a 1.6-mile-long force main to connect the rehabilitated aqueduct to the existing Cedar Creek Water Pollution Control Plant (WPCP) outfall, which discharges and diffuses treated water three miles offshore in the Atlantic Ocean. This paper will review the Design-Build requirements, the chosen innovative solutions during planning for the micro tunneling portion of the project and review the microtunnel construction on the project.

Jun 23, 2024

By James S. Goff, Stephen J. Navin, Wesley W. Moulton

The White River Shale Oil Corporation (WRSOC), as agent for Phillips Petroleum Company, Sohio Shale Oil Company, and Sun Shale Oil Company, is developing a prototype oil shale facility on Federal Lease Tracts Ua and Ub, located in northeastern Utah and known as the White River Shale Project (WRSP). The Ralph M. Parsons Company (Parsons) has been retained as the design engineer/constructor for the project. This paper covers the successful pregrouting of the Bird's Nest Aquifer in conjunction with the sinking of a 9.2-m (30-ft)-diameter shaft and driving of a 3.89-m by 5.64-m (18.5-ft by 12.75-ft) decline. The Bird's Nest Aquifer derives its name from the ellipsidal solution cavities caused by leaching of random-size nacholite nodules. The sodium bicarbonate nodules occur as persistent bands within a matrix of siltstone and marlstone. The team, made up of the owner, engineer/contractor, and subcontractor successfully grouted the shaft down to an inflow of less than 0.06 lps (1 gmp) and the decline down to less than 0.12 lps (2 gpm) by injecting a total of 213.190 kg (470,000 lb) of Celtite Terraseal, 113,400 kg (250,000 lb) of limestone dust, and 41,730 kg (92,000 lb) of cement into the formation.

Jan 1, 1985

By Todd Grounds, William R. Reed, VASU GANGRADE, Michael Shahan, Garron Ross, Kunal Singh

Roof bolter canopy air curtains (CACs) are gaining acceptance as a respirable dust control device that can provide roof bolter operators with protection from overexposure to respirable coal mine dust. Both lab and field studies on the effectiveness of roof bolter CACs have been published. Field studies have shown the effectiveness to be variable. However, in all previous field studies, none has been conducted when the roof bolting machine operates downwind of the continuous miner (CM)—a scenario for which the CAC was designed to provide respirable dust control. This study, performed by researchers from the National Institute for Occupational Safety and Health (NIOSH), was conducted to test a CAC on a roof bolter machine operating downwind of the CM. The results of testing demonstrated that the roof bolter CAC can effectively provide respirable dust protection for roof bolter operators with dust control efficiencies ranging from 11 to 40%.

Oct 14, 2020

By J. Kang

To obtain microbial cultures with the ability to efficiently leach copper ion from chalcopyrite, microbial samples from acid mine drainage sites were mixed and then subjected to mutation using sodium nitrite (NaNO2), diethyl sulfate (DES), ultraviolet light (UV) and their combinations. Mixed cultures with a much stronger ability to leach copper ion (identified as mixed cultures A, B, C, D and E) were obtained by screening on leaching media. Leaching tests showed that mixed culture E had the best leaching performance and increased the Cu2+ content by 120.5% in 20 days, compared to a control without mutation treatment. Mixed culture E grows well on Leathen leaching medium supplemented with Cu2+ (at a final concentration of 32 g/L), reaching 3.2 × 107 cells/L after 15 days, while the unmutated control culture cannot survive beyond 15 days, indicating that mixed culture E is more tolerant of high concentrations of Cu2+. In addition, continuous leaching by mixed culture E is very effective at a pulp density of 1%. These results suggest that it has the potential for industrial applications.

Jan 1, 2008

By J. Duran

Mining at Newmont Mining?s Phoenix?s Mine is currently done by surface operation in the Fortitude and Bonanza pits at a combined rate of approximately 100 kt/d (110,000 stpd). A portion of the orebodies targeted by these pits was mined previously by underground mining methods, dating back to the mid-to late-1800s. Most of the underground workings (voids) are located in the west central area of the Bonanza pit. The surface operation requires excavation through voids, and there is a risk of pit floor subsidence that has to be addressed. This article presents a review of the risk mitigation measures that were implemented to reduce risks to people and equipment working in the area.

By Maurice Grieves

GC. Waterman Mr. Grieves' paper on "Shaft Sinking Today--A Boring Business Tomorrow" in the Dec. 1981 issue of MINING ENGINEERING is an excellent description of recent improvements in speed and costs of shaft sinking. However, shaft boring techniques are not a recent development. In 1935-36, the Idaho Maryland Co. (gold), Grass Valley, CA, bored the vertical 1.5-m-diam Idaho No. 2 ventilation and supply shaft to a depth of 346 m. The equipment was designed and perfected by Branner Newsome and was, I believe, described in "Transactions." Pickands Mather later bored a 1.8-m-diam shaft to a depth of about 305 m using Newsome's design. The Idaho No. 2 shaft cut through hard gabbro, diabase dikes, soft serpentine, very incompetent ankeritized serpentine and a strong fault zone. The completed shaft did not require timber support; the fault zone was cemented. Equipment consisted of a rotating 1.5-m or 3-m core barrel with a slotted bottom which cut a 76-mm kerf. The cutting agent was chilled steel shot introduced into the slot as needed. The driving mechanism and core barrel were lowered into the shaft, the former secured to the wall and the inshaft "engineer" operated the motors which controlled core barrel rpm and advance. After an advance of 1.5-3 m the equipment was lifted out of the shaft, a half(?) stick of powder cut off the drilled core, and it was hoisted to the surface with a cable attached to an eye bolt at the tope of the core. A "shift" consisted of three men: hoistman, equipment operator (down the shaft), and a surface laborer. Advance was variable as equipment and techniques were perfected. Near the end of the job advance was, as I remember, 1.5-3 m per shift. Costs were, as I recall, about $115/m. Up to 4 m cores were lifted out in one piece and swung to the dump by a stiff leg derrick. The Newsome equipment was relatively inexpensive to build and operate and his method should be (more than?) competetive with the methods described by Grieves. W.E. Hawes The author, in mentioning the South African developments of the cactus grab, ignores the parallel development of the Cryderman mucker in this hemisphere. Perhaps a slightly longer history of blind hold drilling would have been in order. Blind shaft boring got a major boost at the Nevada Test Site, as part of the nuclear weapons development, when it was essential to be able to quickly assess depth for weapons testing. One of the early civilian attempts to utilize this technology occurred when Kerr McGee drilled the shaft at the Section 19 Mine at Ambrosia Lake, NM. This twin masted drill rig belonged to a subsidiary of Kerr McGee, not Shaft Drillers. This was a difficult hole, due to bentonitic shales that decomposed. Later advances in drilling fluids eliminated this problem. The Conoco project referred to is not entirely correct, and needs minor amplification: The depth of the development (not pilot) shaft drilled was 684 m, at a diameter of 3.05 m, with a 2.16 m ID hydratatic casing being installed to a depth of 669 m. Conoco Inc., not Challenger, executed the drilling, however Conoco did use Challenger Drilling Company's Rig 14-S. Rig 14-S is unique in that it is specifically designed for shaft drilling, rather than being an oil field rig used for this service. All work was planned and directed by Conoco personnel. Deviation of the aforementioned shaft from vertical was 31 cm. Two smaller shafts (pilot shafts) were drilled with the same rig. These smaller ones were drilled 1.8 m in diameter, to depth of 665 m and cased with .91-m-diam casing, to depth of 650 m to be used as pilot holes for 5.5 m diameter shafts. The total elapsed time for all drilling, casing, cementing, and rig moves was 50 weeks. The constraints mentioned of vertical accuracy, torque, wall collapse, rig set up time, lining time etc., pumping out, etc., may not be as serious as the author implies. First, vertical accuracy: The record achieved by Conoco indicates that with reasonable care, good accuracy can be obtained, primarily by having a fair amount of weight in tension, rather than on the drill bit. Second, the amount of torque that can be applied to the drill stem has been greatly increased by the introduction of larger diameter drill stems, such as the new Hughes rig uses. Third, risk of collapse of shaft walls is minimal with the proper mud program. Basically, the walls are supported by the weight of the mud and it is not removed until some type of casing is installed and grouted in. Here in the real challenge of blind drilling shafts to devise a more economical method of lining bored shafts than using steel casing. The remaining issues were minor at Conoco, provided adequate planning and scheduling takes place. An excellent paper, "Shaft Drilling--Crownpoint Project" by Hassell H. Hunter, presented at the Fifth Uranium Symposium in Albuquerque in Sept. 1981 contains details of this project. A combination of blind boring, with enlargement by either mechanical means (boring machine, or tunneling type shield) or slabbing down, with muck removal through a larger bored shaft would seem to be the future trend in shaft development. The economics, especially for wet conditions, favor this concept over conventional sinking, as has been demonstrated at Kerr McGee --Churchrock, UNC--Churchrock and Conoco at Crownpoint.

Jan 3, 1983

By V. Arslan

The primary advantage of using oxalic acid is in the formation during leaching of soluble complexes that are both photo-chemically and microbiologically degradable. In this study, chemical leaching and bioleaching methods were used for quartz sand. Using the ANOVA-Yates test technique, the effects of solids rate, temperature and oxalic acid rate were investigated using chemical leaching experiments. Solids rate and temperature were effective parameters on the chemical leaching. After the tests, the removal of Fe2O3 increased to 97.02%. Finally, the iron content decreased from 0.315% Fe2O3 to 0.0094% Fe2O3 in the leached quartz sand. The effect of pulp density was investigated on the bioleaching experiments. The bioleaching of iron oxides by Aspergillus niger resulted in removal of 47.70% of the total iron in the quartz sand, and suggested that this strain has not a good ability to remove iron impurities from quartz sand in comparison with chemical leaching. The iron content decreased from 0.315% Fe2O3 to 0.164% Fe2O3 in the bioleached quartz sand.

Jan 1, 2009

By Qinghua Tian, Zhipeng Xu, Dong Li, Xueyi Guo, Liu Zhu

A possible process for extracting tellurium and antimony from a residue generated in alkaline sulfide leaching of telluriumbearing alkaline skimming slag was proposed. The central composite design was employed to define the optimal leaching conditions for tellurium and antimony extractions from the residue sample. Using a 23 factorial matrix, the best performance for tellurium and antimony extractions (96.36% and 99.64%) were obtained at 175 g L−1 Na2S concentration for 80 °C during 60 min. The notable degree for influencing tellurium and antimony extractions was leaching temperature, Na2S concentration, and leaching time. The analysis of variance results indicated that the regression model was significant and suitable to be used for the optimization of independent variables within the range. The model was validated under optimal operating conditions based on the experimental design results. Antimony and tellurium in the leach liquor were separated and recovered by H2O2 oxidation followed by Na2S-Na2SO3 reduction. 98.93% of antimony was precipitated selectively during H2O2 oxidation, while 95.06% of tellurium was precipitated during Na2S-Na2SO3 reduction. NaSb(OH)6 and elemental tellurium was obtained as the product.

By John A. Cordes, Thomas F. Torries

INTRODUCTION Nonfuel mineral markets have always tended to be volatile due to the cyclical nature of demand and the difficulty of fine-tuning capacity and output levels. Experiences over the past two decades, however, have been especially perverse. This short volume attempts to provide insights into this era and some guidance for the future. The papers are linked together by their focus on issues of productive capacity. At the same time, they provide a useful description, explanation and evaluation of the industry's evolution and its implications for the near-term. The regnant pessimism of the 1980s was very much alive during the early months of 1987 when the theme of this symposium was formalized. Indices of real metal prices had only inched-up from all-time lows in 1986. The hemorrhage of corporate red-ink was still a powerful memory. Frustration with the apparent inability of closures and reduced capacity utilization to re-establish market balance was strong. To the more pessimistic, the industry appeared destined to a quarter century of persistent excess capacity and nonremunerative prices. The conditions to support this pessimism were readily apparent. Economic literature tells us that surplus capacity is a complex, usually unexpected and often long-lived phenomenon caused by the interaction of at least three factors. The first is a rising capital-intensity of production. The second is the presence of aggressive new entrants whose capacity additions are typically (1) trade- dedicated because local market absorption is low; (2) cost-competitive for natural or policy-support reasons; (3) heavily burdened by debt-repayment obligations; and (4) viewed as important sources of employment, regional development and foreign exchange. The third, and often decisive, factor is an unanticipated decline or stagnation in demand growth for the sector's output. The first two factors combine to increase competitive pressures and to raise the actual or perceived cost of idle capacity. Market rigidity reduces responsiveness to price signals and distorts the burden of adjustment. Demand atrophy exacerbates these pressures especially on unprotected producers in the higher quadrants of the global supply function. In the mid-1980s, evidence of the existence of all these factors were easy to cite. Culprits were often identified. Claims for protection, global sharing of adjustment burdens, and imposition of disincentives to new capacity regardless of inherent economic status were co-dominant themes with the buzz-words of "competitiveness" and "level playing fields." This, notwithstanding the suggestions of the same economic literature that excess capacity persists longest when interventionist remedies permit it to continue. Today, only eighteen months later, the mood is decidedly more buoyant and most analysts are cautiously optimistic about the next few years. Many informed analysts have concluded that the excess capacity of the recent past has disappeared. For the present, the marketplace seems to have restored a balance between supply and demand. That the adjustment took five or six years longer than anticipated attests to the dramatic structural changes which have characterized the industry. The implications of these changes for the future are unclear. Were the macroeconomic shocks, coupled with industry transitions, so extraordinary in magnitude and timing that experiences were unique? If so, the next years may hold more stable and cautious times for the industry. Alternatively, is the inherently volatile metals industry facing a future of cycles equally perverse in terms of amplitude and duration due to economic uncertainities and inflexible industry response capabilities?

Jan 1, 1989

By N. S. Tanious

This work was completed prior to the author's appointment with the Nuclear Regulatory Commission (NRC). Therefore the views and conclusions contained in this paper are those of the author and should not be interpreted as necessarily representing the official policies or recommendations of the NRC. Reference to specific brands, instruments, or trade names is made to facilitate understanding and does not imply endorsement by the author or the NRC.

Jan 1, 1987

By R. R. Klimpel

The industrial processing of ores containing precious metals such as gold is generally very mining site specific, e. g. McQuiston and Shoemaker (1985), vanZyl et. a1. (1988), Fuerstenau and Hendrix (1990), and Hausen et. a1. (1990). The variety of various mineral processing techniques employed is extensive and includes size reduction, size classification, gravity separation, roasting, cyanidation, chemical and/or bacterial leaching, ion exchange and/or solvent extraction, and froth flotation. Gold occurs in the earth's crust on the average in very low concentration estimated by various sources to be between two and five parts per billion. For a specific ore deposit to be considered as an economic source of gold, generally a concentration of near one part-per million (one gram / Mton) or more is required. From a mineralogical viewpoint, gold often occurs in its native or free form. However, it can also occur as gold compounds of telluride, antimony, and selenium contained within either sulfide minerals such as pyrite, arsenopyrite. and pyrrhotite or, to a less extent, within silicate, oxide, and carbonate minerals. In many locations, alloying of gold with silver is found. The optimal processing method for recovering gold is determined by many factors such as the mineralogy of the gold bearing minerals as well as the associated mineralogy of the gangue minerals, the type of liberation profile of the gold bearing minerals upon' size reduction, and the particle size of the gold itself. Free or native gold particles can vary in size from the extreme of large nuggets to very finely disseminated gold associated in a complex sulfide mineral matrix. Typically free gold particles of several hundred microns or larger are efficiently recovered by one or more of a number of gravity based techniques due to the large specific gravity differential of gold (19.3 to 15) versus typical gangue minerals (5 to 2.7). When gold is associated with a sulfide (usually pyrite) matrix, historically most industrial plants involving size reduction and subsequent liberation have operated in a size range of 50 to 80 percent less than 75 microns. Grinding finer than this has been normally not economical unless the ore has a high gold content in the feed ore.

Jan 1, 1998

By W. Lorenz

The responsible treatment of natural resources, including mineral raw materials, contributes to the protection of the environment. The management of resources and the environment encompasses reduction of emission through application of environmentally friendly processes, economical use of resources, recycling and substitution, application of energy saving processes, land reclamation and recultivation, as well as a well balanced regional and land use planning. A variety of materials are needed for measures to protect the environment. Many of these materials are, or to a large percentage consist of, mineral raw materials, which become effective either directly or indirectly in their natural state or after an appropriate industrial treatment. We call them environmental minerals. Although extraction and processing of these mineral raw materials destined for use in the protection of the environment, may have a negative impact on the environment on one hand, preventive or curing measures to protect the environment on the other would not be possible without them. We have to accept certain smaller burdens of the environment, which naturally will have to be kept at the lowest possible level, in order to avoid or minimize greater damages to nature. In the broad group of industrial minerals and rocks there is a wide variety of mineral raw materials, which lend themselves for countermeasures in favor of the environment. To hamper their production and processing through excessive environmental regulations would make the protection of the very environment more difficult or even impossible.

Jan 1, 1995

By T. Brozell

"Communities near sand mines and processing facilities in Wisconsin have raised concerns regarding possible exposure to ambient respirable (PM4) crystalline silica. Fenceline PM4 (particulate matter captured with a 50 percent cut size of 4 micrometers aerodynamic diameter) crystalline silica concentration data were compiled at Fairmount Santrol Inc., Mathy Construction Inc. and U.S. Silica facilities. The long-term average PM4 crystalline silica concentrations ranged from 0.05 to 0.45 micrograms/m3 when sample values below the limit of quantification were treated as zeros. All of the long-term average concentrations were less than 20 percent of the Minnesota Department of Health’s and California Office of Environmental Health Hazard Assessment’s PM4 crystalline silica chronic reference exposure level of 3.0 micrograms/m3. The measured concentrations were consistently within the local background concentration range.Air Control Techniques P.C. compiled the sampling data using Thermo Fisher Scientific Inc. Partisol 2000i PM2.5 samplers meeting the requirements of 40 CFR Part 50, Appendix L. These samplers were adapted to capture particulate matter with a 50 percent cut size of 4 micrometers. The use of polyvinyl chloride filters allowed for crystalline silica X-ray diffraction analysis by NIOSH Method 7500. The combined use of adjusted PM2.5 filter-based samplers with Method 7500 crystalline silica analyses provided for a highly sensitive limit of quantification of 0.31 micrograms/m3 of PM4 crystalline silica.BackgroundIn 2005, the California Office of Environmental Health Hazard Assessment (OEHHA, 2005) adopted a 70-year chronic reference exposure level (REL) for PM4 (particulate matter having aerodynamic diameter of 4 micrometers or less) crystalline silica. The REL was extrapolated from occupational hygiene epidemiological studies, essentially all of which were based on in-plant PM4 crystalline silica measurements. No well-established ambient sampling technique for PM4 crystalline silica was available at the time that OEHHA adopted its REL. The personal samplers used for the National Institute for Occupational Safety and Health (NIOSH) Method 0600 in-plant sampling of employee exposure were designed for concentrations higher than ambient levels. The limit of quantification (LOQ) of PM4 crystalline silica on NIOSH 0600 filters was too high for the effective sampling of ambient air."

Jan 1, 2015

By D. H. Horswill

The Red Dog Mine is the world’s largest zinc mine. It is located 145 km (90 miles) north of the Arctic Circle in northwest Alaska on lands owned by the Inupiat Eskimo organization called the Northwest Alaskan Native Association (NANA) Development Corp. NANA acquired surface and subsurface title to this resource as part of its settlement lands under the Alaska Native Claims Settlement Act. In 1982, NANA entered into a lease agreement with Cominco. Cominco is an international mining company focused on zinc. It built and operates the Red Dog Mine. NANA’s principle objective in the development of the mine was to have it contribute to the development of sustainable communities within their traditional region. This meant creating job opportunities in a modern economy while preserving the traditional subsistence lifestyles. Preservation of their culture, language and traditions are as important to the NANA people today as they were in the past.

Jan 1, 2000

By D. Malhotra

Metallurgists are continuously striving to improve recovery. But optimization of profits and/or revenue dictates that recovery can be sacrificed at the expense of throughput. However, recently there has been significant interest in the concept of retreatment of tailings to enhance recovery and hence revenue while maintaining high throughput. The authors discuss the technical and economic factors that need to be considered in evaluation of the feasibility of retreatment of tailings including the potential feed streams (i.e., rougher tailings, cleaner tailings, etc.) and processing options (i.e., physical beneficiation, chemical processing, etc.). A case study for base metals is discussed.

Jan 1, 1994

By Donald O. Rausch, William E. Hawkes, Odin Jr. Strandberg

INTRODUCTION Kennecott Copper Corp.'s Tintic Div. Burgin mine is a 725 t/d (800 stpd) lead-zinc producer, located in the east Tintic mining district, 5.6 km (3.5 miles) east of Eureka, and 130 km (80 miles) south of Salt Lake City, UT. The Burgin mining complex includes three shafts, shop, dry, and office facilities, and a 454 t/d (500 stpd) flotation concentrator. Production and ser¬vice entry to the mine is through a 4.26-m (14-ft) diam concrete-lined production shaft, and a three-compart¬ment rectangular shaft. A connection with the Apex No. 2 shaft and the Burgin mine provides a second exhaust airway. Bear Creek Mining Co., Kennecott's exploration sub¬sidiary, began exploration of the Burgin mine area in 1957 with the sinking of the 335 m (1100 ft) deep No. 1 shaft. A station was cut at the 320-m (1050-ft) level of the shaft and extensive drifting, diamond drilling, and some winzing and raising confirmed the presence of a major lead-zinc ore body, the main portion of which oc¬curred below a hot saline water table. Surface facilities and the 4.26-m (14-ft) diam concrete-lined production shaft were completed in 1966. Ore production has been steadily increased to 725 t/ d (800 stpd) since 1966. The collar elevation of the No. 2 production shaft is 1731 m (5678 ft) above sea level. The mine's three levels, the 1050, 1200, and 1300 are at elevations of 1414.6, 1356.4, and 1338.1 m (4587, 4450, and 4390 ft), respectively. The approximate elevation of the hot saline water table is 1382.6 m (4536 ft). The earliest production on the east Tintic mining district was from the Eureka Lilly mine in 1909. Later, in 1916, the Tintic Standard mine was discovered, and pro¬duced ores with a gross value of $80,000,000 through 1956. The Eureka Standard mine located 1.6 km (1 mile) southwest of the Burgin mine produced ore worth $12,000,000 from fissures in the Tintic quartzite along the Eureka Standard fault. Ores produced from the east Tintic mining district through 1971 are valued at $151.3 million. The production from the entire Tintic district is approximately $500 million. GEOLOGY The Burgin ore deposit occurs as replacement bodies in limestone, dolomite, and shale. Numerous fault struc¬tures and rock formations make the mine geology com¬plex. The mineralization is strongly associated with sev¬eral major geological features, which include the Ophir formation, consisting of limestone, dolomite, and shale as the favorable host rock, and the structural influence of the east Tintic thrust, the Eureka Standard and sev¬eral other significant faults. Fig. 1 depicts the Burgin 1300 level workings, a generalized configuration of the ore body, and the main geological features. Covering the quartzite, limestone, dolomite, and shale facies enclosing the Burgin ore deposit are 183 to 274 m (600 to 900 ft) of andesite. The Burgin ore consists of massive sulfide and oxide lead-zinc ores with significant silver and cadmium val¬ues. The main ore minerals are galena, cerrusite, and spalerite. Extensive manganese oxide deposits are associated with the lead-zinc ore occurrence. A limited

Jan 1, 1982

By N. Cruickshank

Gas outbursts occur as sudden eruptions of gas and minerals from the face of a mine heading as the heading approaches potentially burst-prone (or anomalous) zones of rock. Such outbursts have resulted in considerable damage to equipment and in some cases lives have been lost. Characterizing these anomalous zones and detecting their presence is therefore vital to safe mining. Certain significant characteristics associated with the outburst-prone zones are: (1) the presence of fluid inclusions, (2) unstable stress conditions, (3) characteristic geologic structures, (4) certain physical properties of the rock, and (5) micro-seismic activity. Geophysical and engineering methods for detecting these characterisitcs are discussed and are considered with respect to their past success and future potential for predicting gas outbursts.

Jan 1, 1985

By Werner Wenzel

Whisker growth during the reduction of iron ores The phenomenon of whisker growth can be observed in a number of solid-state crystallisation reactions, which, apparently, appear in6 the first instance as though they are two independent processes. From such reactions, one can find the whisker type of crystallisation on solid bodies. These examples show, that the whisker growth is neither due to certain crystal forms with predetermined temperature and pressureranges, nor to parameters, which, ad hoc, are necessary for the crystallisation. In all cases the whisker formation seems to depend on nucleation and the availability of crystallizing material in direct environment of such a nucleus, under given parameters of a system. One can generally state, that the research of whisker growth with the aim of establishing proper physical laws, to determine the foundamentals of whisker formation and whisker growth, is at present under active study. We can conclude that, although a great number of hypothesis have been proposed; no concluding theory has yet been arrived at. In the institute for ferrous metallurgy of the technical university of Aachen, research works are undertaken to find an explination to various metallurgical phenomena through help of whisker growth which was observed under varying conditions of iron ore reduction. This was possible by using refined microscopic techniques through which the whisker growth on solid oxyde substrates is detected. This refined technique combines the latest methods of surface examination with the help of stereo-scan-electron-microscope and proper and carefull preparation of surface to be stereo-scanned.

Jan 1, 1971

By John W. Trammell

The White Pine Mine is situated in Ontonagon County, approximately six miles south of Lake Superior, in the Upper Peninsula of Michigan (Figs. 1 and 2). The ore body mined by White Pine Copper Company occurs in the lower - most 20 to 25 feet of the Nonesuch formation, a series of middle - to upper Keweenawan shales, siltstones, and sandstones. This paper elaborates on the solution of problems in ore grade control and mine planning arising from the presence of an essentially barren sandstone stratum within the ore body.

Jan 1, 1964