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By Anthony Delvescovo, Gregory F. Aluce, Joseph A. Sopko, Lyle C. Smith

Construction of a 13m shaft through approximately 100m of water bearing soils and another 100m of bedrock was attempted using the Slurry Diaphragm Wall Method with a combination of deep dewatering wells. During excavation within the slurry wall, unexpected groundwater inflows occurred in a structure of loose silty sands. Artificial ground freezing was selected as a remedial method to seal the leaking slurry wall, and as a primary excavation support and groundwater control method for the 30m of overburden material below the base of the wall. Design of the ground freezing system required analysis of both the structural requirements of the unsupported soil, the frost effects against the existing slurry wall and the groundwater control in the area of the breech in the wall. These design procedures are summarized, and construction and excavation techniques and corresponding schedules compared in this paper. Advanced drilling techniques, ground freezing quality control and general contractor procedures for efficient shaft construction are discussed.

Jan 1, 1993

By R. L. Keiski, H. Miettunen

"The serpentinite massifs of Hitura are located in the municipality of Nivala in Central Finland. The most important sulfide minerals in the ore are pentlandite, pyrrhotite and chalcopyrite. The ore body consists of amphibolite rocks in contact with serpentinite against mica gneiss. The mean Ni grade of the ore is about 0.7%, and the highest Ni grades have been encountered in close contact. In general, the inner parts of the ore are strongly serpentinized ultramafite and some sulfides have been converted to mackinawite and valleriite. The degree of alteration varies considerably within the ore body and, correspondingly, the grain size of the sulfides varies as well. In this study, flotation tests were done with samples from different zones of the ore body. The flotation test results on the bench scale and plant scale are discussed in terms of the mineralogical features of the ore. Ores in the North Hitura zone differ from those in the South Hitura zone.IntroductionThe Hitura nickel mine, like the entire Nivala area, is part of the Bothnian Svecokarelian schist zone, where numerous Ni-Cu boulders have been discovered since the early 1930s. The Nivala area has therefore been the subject of exploration activities by the Geological Survey of Finland (GSF) and mining companies. The first period of exploration started in 1936, and the Makola nickel deposit, located 3 km south of the Hitura mine, was discovered by GSF in 1937. About 415,000 tonnes of ore were mined at Makola between 1941 and 1954. In 1937, during the exploration and development stage of the Makola mine, boulder tracing and a geophysical survey revealed a significant ultramafic body with some lowgrade Ni-Cu occurrences in the South Hitura zone.In 1961, GSF continued the exploration program in the area and discovered the North Hitura ultramafic complex and its allied Ni-Cu occurrences. In 1964, Outokumpu Oy bought the property and, after underground exploration, comprehensive mineralogical studies, process research and a feasibility study, production started in 1970 (Isohanni et al., 1985)."

Jan 1, 2015

By T. Sun, J. Kou, D. Tao

"The reduction behavior of low-grade lump hematite ore (100% 20-50 mm) has been evaluated by performing a series of coal-based direct reduction roasting tests followed by magnetic separation under different process conditions such as roasting time, temperature, reductant type and dosage, and grinding stage. To investigate mineralogical transformation and reduction characteristics of lump hematite ore, roasted ores and magnetic concentrates were analyzed by X-ray diffraction (XRD), scanning electron microscopy (SEM) with energy dispersive spectrometer (EDS) and polarizing microscope. The results have shown that a magnetic concentrate with a total iron (TFe) grade of 90.7% was produced at an iron recovery of 83.9% when lump hematite ore was reduced at 1,200° C for five hours using 35% lignite as a reductant followed by two stages of grinding, for 20 min and 30 min each, and low-intensity magnetic separation. The degree of reduction is closely related to the change in the internal structure of the roasted ore. Lignite with a higher fixed carbon content and lower ash generated roasted ore with a more highly porous structure, which provides better permeability of reducing gas. Magnetite, fayalite and wustite were the dominant intermediates of coal-based direct reduction of lump hematite ore. Parts of wustite reacted with SiO2 or Al2O3 in gangue and generated fayalite or hercynite. FeO in fayalite was replaced by CaO to produce wustite, which was finally reduced to metallic iron. However, Fe in fayalite and hercynite produced with low contents of CaO cannot be reduced into metallic iron. Introduction Low-grade hematite ore, characterized by low Fe content and high SiO2 content, cannot be treated by traditional mineral processing methods due to its nonmagnetic property and fine liberation size. The direct reduction (DR) roasting process has been studied and proven satisfactory in treating difficult-toprocess ores, such as low-grade nickel laterite oxide ore (O’Connor et al., 2006; Cao et al., 2010; Guo et al., 2011), high phosphorus oolitic hematite ore (Li et al., 2012; Yu et al., 2013), low-grade hematite ore (Wei et al., 2013;) and iron ore tailings (Li et al., 2010), which greatly expand the resources of usable iron ores (Srivastava and Kawatra, 2009). Therefore, the use of a DR process to treat low-grade hematite ore to produce direct reduced iron (DRI) is of great interest to many researchers of mineral processing and metallurgy (Weissberger et al., 1986)."

Jan 1, 2015

By X. Li, J. C. Stankus, J. Ma, H. Chen

Accurate evaluation of stress and geological conditions is critical to ground control design in underground openings. For a slope entry, the problem is more difficult, because a slope will transverse through many types of strata. Over the last few years, Keystone Mining Services LLC, an affiliate company of the Jennmar Corporation, has developed a new methodology designated the “Stress, geologic and support design system (SGS).” The method analyzes the stress, geology and support design for a mine slope. The method has been successfully applied in various new slopes across the United States, and ground control plans developed using SGS methodology has helped coal companies successfully pass review and/or hearings by state and federal regulatory agencies. With a case application in the U.S. Midwest, this paper details and demonstrates the main concepts of the methodology, including (1) analysis of stress and geology and identification of strong, fair and weak zones along the slope; (2) primary and supplemental support design; (3) long-term support design in the form of steel square sets or arches and (4) validation of the support design utilizing advanced finite element computer modeling.

Jan 1, 2011

By D. Spitty

Mining companies often have many disparate systems and repositories of data in their organization for planning and scheduling outbound mining supply chains. To optimize ore quality, throughput, transport logistics scheduling and delivery to contract, planning teams need to model and manage the value chain accurately and achieve a consolidated, enterprise view of operational performance. Planning teams coordinate a complex series of activities to ensure production forecast accuracy. They also negotiate and execute complex contracts often with disconnected and limited visibility of production forecasts, quality, and variability with little ability to accurately market potential product specifications over longer time horizons. This market is becoming increasingly dynamic, with increasing variability and the time for planning teams to react is decreasing. This is driving the need towards using science to generate the necessary scenarios required to support the decisions made by planners on a frequent basis. Resource-to-Market is a highly flexible, user-configurable supply chain management system for the mining and metal processing industries. This provides an integrated view of the supply chain that supports the dynamic nature of planning in a mining company, embraces variability of elements across the supply chain with the ability to optimize across all the functional areas of the operation to enable optimal decision making for often competing objectives such as throughput, quality, cost, revenue and profit.

Feb 23, 2014

By J. Engel

The talc deposit in Gemerská Poloma lies 350 to 400 m under the surface; the whole deposit is located at the highest-level drinking-water-source protection zone, therefore, opening in this mining area (a well, a ramp) requires higher investment, operational and water protection costs. The adit, outside the mining area, with a length of 4200 m, rising with a gradient of 3? and a diameter of 10.3 m2, allows free passage/runoff of mining-water through the adit, and guarantees low costs for transporting produced material and equipment. This solution is in total less demanding from both investment and operational point of view.

Jan 1, 2007

By Ross A. Scarano

It wasn't until late 1976 that the U. S. Nuclear Regulatory Commission (NRC) set about to address the uranium tailings issue in earnest. It was decided that a generic environmental impact statement related to uranium milling operations would be prepared to support formal tailings regulation. We recognized, however, that this was a task that would take many years to complete and we could not take a "business as usual" approach to tailings management in the interim; the issue was too sensitive. To facilitate licensing during the interim period, the NRC developed guidelines in May 1977, for the industry in the form of the following performance objectives for tailings management: Siting and Design 1. Locate the tailings isolation area remote from people such that population exposures would be reduced to the maximum extent reasonably achievable. 2. Locate the tailings isolation area such that disruption and dispersion by natural forces is eliminated or reduced to the maximum extent reasonably achievable. 3. Design the isolation area such that seepage of toxic materials into the groundwater system would be eliminated or reduced to the maximum extent reasonably achievable. During Operations 4. Eliminate the blowing of tailings to unrestricted areas during normal operating conditions. Post Reclamation 5. Reduce direct gamma radiation from the impoundment area to essentially background. 6. Reduce the radon emanation rate from the impoundment area to about twice the emanation rate in the surrounding environs. 7. Eliminate the need for an ongoing monitoring and maintenance program following successful reclamation plan. 8. Provide surety arrangements to assure that sufficient funds are available to complete the full reclamation plan.

Jan 1, 1980

By Brian T. Field

Development of cost effective technology for whole ore roasting of refractory gold ores is an important requirement to increase recovery rates as ore grades become more carbonaceous and sulfidic in nature. It is important that the industry focus on the further development of cost effective, environmentally sound technologies to roast this ore. Traditional technologies have centered around bubbling and circulating fluidized beds with and without oxygen enrichment. Fuller Company has developed a process to flash roast refractory ores in its pilot plant with the R&D results, commercial scale up issues, and indicative economics discussed herein.

Jan 1, 1993

By Saul J. Esdalera

The flotation response to amine collector and its adsorption on the mineral surface was studied as a function of pH and collector concentration. The flotation response increased in the order anorthite + intermediate plagih clases + albite, while adsorption of amine obeyed the Freundlich equation r = A.Cn. This adsorption is proposg to be due to a covalent bonding between the AlOH surface pups of the plagioclase altered outer layer and the amino group of the collector. Finally, the study showed thatmtial dissolution of the solid phase provides the necessary conditions for amine attachment on the mineral particles and consequent flotation.

Jan 1, 1980

By Frank E. McCall, Michael A. Trevits, Fred Garcia

The U.S. Bureau of Mines investigated the influence of atmospheric pressure changes on methane gas migration through mine seals at a mine site located in the Pittsburgh Coalbed. The mine gained access to a coal reserve through part of an abandoned mine and constructed nine seals to isolate the extensive old workings from the active mine area. Underground problems were experienced when atmospheric pressure fell, causing methane gas to migrate around the seals and into the active workings. During mining operations, methane gas levels exceeded legal limits and coal production was halted until the ventilation system could be improved. When mining resumed with increased air flow, methane gas concentrations occasionally exceeded the legal limits and production had to be halted until the methane level fell within the mandated limit. To assist the ventilation system, a pressure relief borehole located in the abandoned workings near the mine seals was proposed. Preliminary estimates by a gob gas simulator (computer model) suggested that a 0.76 m (2.5 ft) diameter pressure relief borehole with an exhaust fan would be necessary to remove enough methane from the abandoned area so that the ventilation system could dilute the gas in the active workings. However, by monitoring methane gas emissions and seal pressure, during periods of low atmospheric pressure, the amount of methane gas that migrated into the active mine workings was calculated. Researchers then determined that a relief borehole, 20.3 cm (8-in) with an exhaust fan could remove at least twice the maximum measured volume of migrating methane gas. Because gas concentrations in the abandoned workings could potentially reach explosive limits, it was proposed that the mine eliminate the exhaust fan. Installation of the recommended borehole and enlarging two other ventilation boreholes located in the abandoned area reduced methane gas leakage through the seals by at least 63%.

Jan 1, 1995

By J. W. Markham, M. Atiemo, K. Yoshida

INTRODUCTION In open-pit mining of high grade ores, operators of mining equipment in a pressurized cab are protected from inhaling radioactive aerosols by the use of filtered air through a pre-impactor and a HEPA (high efficiency particulate air) filter. At present, a limited amount of scientific information is available on the worker exposure to airborne alpha emitters in the mining of high-grade ores. (Miller, 1977, Nielson, et al, 1979). Factors affecting the worker exposure to alpha radiation are: the grade of uranium containing ores, the environmental conditions inside the mine pit, the protection factor of the cab filter system, the workers' attitude towards the hazards of radioactivity and the adopted contamination control in the cabs. Four series of field samplings were conducted during the fall of 1980 and the summer of 1981 at a mining site in northern Saskatchewan to investigate the extent of protection an equipment operator has against cab internal exposure to airborne alpha emitters with short and long half-life radionuclides. The variation in the breathing zone concentrationnof aerosols as a result of resuspension is discussed, however no attempt was made to isolate and investigate the factors and the associated parameters of the aerosol resuspension phenomenon. AEROSOL RESUSPENSION IN A CAB Mechanisms of Resuspension [Sources]: The presence of radioactive aerosols inside a cab and in the breathing zone of an operator may be due to initial concentration inside [(CO)], concentration in the filtered (HEPA) air (CC, at the flow rate [QC]), the concentration of the aerosol through resuspension ([CR]), the aerosol concentration in the air outside the cab ([CE]) and the concentration of the aerosol in the leaking air [(CL)], at the flow rate of [(QL)]. [Volumetric Equilibrium]: Assuming a uniform concentration of the aerosol in the cab, the aerosol concentration in the breathing zone[ (CBZ)] may be expressed as, [CBZ+CO+CC+CR-CL] If [QC] remains greater than [QL], there is pressurization in the cab. There is particle loss as a result of the leaking air from the tight cab, however the condition of adequate pressurization ensures that [CL ~ CC]. Also for dust particles filtered through HEPA [CC – 0] thus [CBZ + CO + CR] For the radioactive gases, [CBZ - CO + CC] Over long periods,[ CBZ – CE] but if [CBZ] exceeds the value of CE then the aerosol build-up in the cab has, occurred. [Resuspension]: Turbulent bursts are generated within the core region of a cab confinement, such that the down sweeps and the up sweeps of the air bursts cause resuspension of dust from the surfaces. Particles initiate resuspension by four major modes: namely, sliding, rocking, rolling, bouncing or sudden ejection or by combinations of these. Nevertheless the adhesion force between the particles and the surface must be overcome by ejection forces before ejection takes place (Kline et al, 1967, Corino and Bradkey 1969, Kim et al 1971). The phenomenon of particle resuspension from a surface may be expressed in a generalized form as: CR =f (Uz . R . D . H) where CR = Rate of particle resuspension UZ = Surface shear velocity R = Surface roughness parameter D = Aerodynamic diameter of particle H = Frequency of surface airburst A detailed discussion of the individual parameters was attempted for the case of farm tractor cabs by Atiemo (1981), such a treatment is out of the scope of this study. Only the effects of the properties of surface roughness will be discussed. Effect of Surface Properties The roughness of the surface from which dust particles are ejected plays a significant role in the deposition and resuspension of the dust. The roughness or the amount of protrusions greatly influence the intensity of air turbulence in the immediate vicinity of the surface (Schmel 1976). For most roughened surfaces therefore, larger particles may be dislodged in greater quantities than smaller particles. The deposition and the resuspension of dust particles from different surface materials of glass, cloth, metal and rubber have been studied in a wind tunnel (Atiemo, 1981). Field Conditions Mining equipments used are: a backhoe, a bulldozer with ripper attachment and two dumptrucks for high-grade ore operations; a payloader and two dumptrucks for waste handling; and a jumbo drill which drills dynamite holes for ore blasting. All this

Jan 1, 1981

By H. E. Cross

In a minerals-recovery context, the object of exercising control over milling operations must be threefold, 1) To achieve maximum recovery at 2) minimum operating costs and 3) making maximum use of installed equipment, Milling is one of a sequence of operations which have as their objective the economic recovery of minerals from mined ores, All operations in the sequence are interdependent and each is both influenced by and itself affects all other unit processes. The part which grinding plays in achieving each of the above objectives is indicated in the following analysis: In South Africa, the most important mining operations are still the gold mines, A vast reservoir of knowledge and experience in treating the Witwatersrand and associated ores has been accumulated. and recoveries are consistently high, However residue values do vary and frequently this is due to fluctuations in the degree of grind, The significance of changes in residue values is indicated by the following example: On a modern mine, the residue values varied from 0,25 to 0,33 dwi; per 2000 lb, during a period of 6 months, The cash value of the gold thus lost varied therefore between $0.44 and $0.58 per 2000 lb, ton. A residue which assays 0.30 dwt per ton would contain approximately 0.26 dwt as undissolved gold and 0,040 dwt per ton as dissolved gold present in residual solution after filtration. This dissolved gold loss is a function of efficiency of filtration and varies from 0.030 to 0,.070 dwt per ton of residue.

Jan 1, 1970

By Y. A. Attia

A study was performed on selective flocculation and dispersion for ultrafine coal cleaning. A totally hydrophobic polymer, FR-7A, was employed as the selective flocculent for coal, sodium metaphosphate (SMP) as the general dispersant for shale and a synthetic polymeric product, polyacrylic acid-xanthate (PAAX), as the selective dispersant for pyrite. In the adsorption tests, nephelometry and spectrophotometry were employed for determining the concentration of FR-7A and PAAX in the solutions respectively. The adsorption tests indicated that: a) FR-7A has a higher affinity toward coal and pyrite than that for shale; and b) PAAX has a much higher affinity toward pyrite than toward coal and shale. In the selective flocculation tests, the ash-forming minerals can be efficiently rejected using sedimentation as the separation method. However, due to their relatively high specific gravity, pyrite particles co-settle with the coal flocs. To reject fine pyrite, a froth flotation was used for floc separation. The one-step flocculation-flotation process could reject more than 70% of each of the pyritic sulfur and ash minerals, while recovering more than 85% of the coal.

Jan 1, 1990

By W. L. Dalke

The Minntac complex is a facility for production of iron ore pellets located on the Mesabi Range at mountain iron, Minnesota. It consists basically of a mine, crushing plant, concentrator, and a pelletizer producing twelve and one-half million tons of pellets annually. This paper deals with some of the aspects of the grinding circuits; but prior to that discussion, it might be well to briefly describe the flow of material in the operation. The crude ore, containing from 20% to 25% magnetic iron, is mined by 14-yard electric shovels. .the blasted ore is loaded directly into 50-yard side dump railroad cars and hauled to the crushing plant. Crushing is done in three stages, originally, two 60" primary crushers; four 7-foot standard secondary crushers; and six 7-foot short head tertiary crushers were installed, when the plant was expanded to double its capacity, five secondary and ten tertiary crushers were added. These were duplicates of the original 7-foot standard and short head crushers used in the original construction, there was no expansion of the primary crushers but a one-half-million-ton surge pile of tertiary crusher feed was added.

Jan 1, 1976

By G. Vijaya Kumar, B. Srinivasa Rao, G. V. Krishna Rao

The paper outlines the ventilation practices prevailing in coal mines of Singareni Collieries Company (South India). The high annual average surface temperature and relatively steep geothermic gradient results in virgin strata temperatures in the region of 38- 42°C creating environmental problems in the working places. Expensive methods of air cooling can not be employed as the mines are small, and production and productivity do not justify the costs of artificial cooling of mine air. Controlled recirculation of mine air -- the unconventional technique being tried with success in some coal mines in U.K. has been analysed for improving the environmental conditions. The analysis shows that the method is safe, cost effective and can be applied with very little extra equipment.

Jan 1, 1991

By Michael D. Sopko

This paper describes the project implementation from financing, through the engineering and construction phases, mine and process plant commissioning as well as operating experience realized to-date. The single line flowsheet used to process the lateritic nickel ore to a 75 percent nickel matte is presented with emphasis on the main pyrometallurgical operations. These include a rotary dryer, a unique reduction kiln, a 45 000 kVA electric furnace and two Pierce-Smith type converters. Associated infrastructure facilities and services necessary to sustain the 11 300 tonnes (12 500 short tons) per year nickel production rate are also described.

Jan 1, 1979

By F. J. Klingler, K. M. Swaffar, M. M. Rabbaig

The Detroit Water and Sewerage Department has undertaken a $1.2 billion effort to capture and treat combined sewage overflows from various outfalls along the Detroit and Rouge Rivers in Detroit, Michigan. This program is planned to include a 455,000 m3 (120 million gallon) storage and treatment tunnel which will capture overflows from 19 outfalls along the Rouge River. The tunnel is currently in the conceptual design phase, with the bid date planned for mid-year in 2001,and completion planned for August 2005. The total design and construction cost is currently estimated to be approximately $450 million. Current concepts are for the storage and transport tunnel will be 6.1 m (20feet) in finished diameter, with a total alignment length of approximately 12,200 m(40,000 feet). The main tunnel will extend through the Traverse and Antrim Shalerock formations, while approximately 5,500 m (18,000 feet) of various diameter connecting tunnel will extend through both rock and soft ground. The project will include a total of 10 drop shafts, 13 riser shafts, various connecting structures, and a 455,000 m3/day (20 mgd) pumping station. This paper will present a discussion of the design philosophy, subsurface conditions, the various planned components, and the expected design and construction considerations for the proposed Upper Rouge Combined Sewage Overflow Tunnel.

Jan 1, 1999

By James A. Finch, Xiang Zhou, Yue Hua Tan

Methods to determine frother functions, control of bubble size and froth properties, are now widely used. Some collectors also exhibit frother functions which are less understood. Using a water-air system in a mini-mechanical flotation cell setup, this paper determines bubble size and water overflow rate for three amine collectors: one primary amine (dodecylamine, DDA) and two commercial ether amines (Flotigam® EDA and Flotigam® 2835-2L) and their combination with two common frothers, MIBC (methyl isobutyl carbinol) and PPG 425 (polypropylene glycol, molecule weight 425). Compared to the frothers, the amines were less effective in reducing bubble size, giving larger minimum size and the two commercial amines showed evidence of coalescence at low concentration. In blends, at fixed frother dosage, frother eliminated the coalescence but as amine concentration increased the amines dominated bubble size. Water overflow was a strong function of reagent type. For fixed 1-cm froth depth, PPG 425, Flotigam® 2835-2L and to a lesser extent DDA produced overflow while MIBC and Flotigam® EDA did not. In blends with frother overflow increased except with Flotigam® EDA. Mechanisms are briefly explored. The principal benefit of blending identified is the elimination of coalescence if residual concentration of the two commercial amines is below ca. 10 ppm.

By Jim Morris, Davin Ishii

The major source of water for Southern Nevada is Lake Mead. However, the existing pump station located on Lake Mead is inadequate for one of the fastest growing areas in the United States. In order to keep up with the increasing demand for water in Las Vegas, a parallel system is being constructed. The intake for this system is being constructed approximately 68.58m (225ft) below the surface of Lake Mead. The lake tap consists of installing casing and secant piles to support the shaft collar, excavating a 21.34m (70ft) deep intake shaft, and installing a3.66m (12ft) diameter steel liner.

Jan 1, 1999

By F. G. Miller

To improve coal flotation Bethlehem Steel Corporation's Research Department has been working on a number of related problems based on both physical and chemical factors. The ideal is of course full understanding of the interrelated effects of both sets of factors. However, detailed studies5 of each physical factor have produced immediately usable results in commercial coal flotation and are providing the bas is for our current work on the effect of chemical factors. Since these detailed studies provided the basis for the present investigation of froth sprinkling, their key results are highlighted here. For example, for a highly floatable bituminous coal, particle floatability was shown to be directly related to the percentage of coal surface available for bubble attachment and inversely related to particle mass. The data also underscored that in a given size range the addition of mineral matter to a particle makes flotation more difficult both by adding to the mass of the particle and by decreasing the amount of coal surface available for bubble attachment. Size also affects particle floatability in that as the size of particles of a given specific gravity increases, mass increases and the ratio between coal surface area and mass of the particle decreases. Flotation tests showed that any recovery of the coarsest particles in a coal feed is accompanied by progressively higher recovery of finer and finer particles. Consequently, fine high-specific-gravity particles are recovered with the coarse good coal. As a result, when the entire 14M x 0 size range is processed by froth flotation, the good coarse coal is contaminated with high-specific-gravity fines. It is the small quantities of these fines that must be rejected from the froth product if flotation efficiency is to be increased above its present level.

Jan 1, 1968