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By H. F. Haller

Large volume sand plugs are placed in tunnels to stem underground explosions at the Atomic Energy Commision's Nevada Test Site. Placement through long, narrow accessways is like stemming coyote holes and blast chambers in mining and construction. Experience with several other methods showed that pipeline conveying could be preferable. Small scale tests justified predicting a placement rate of 20 y. p. h. at 1000 feet, with a modified gunite placer. The initial production trial placed over 2000 c. y. and was estimated to effect a small cost saving after paying for new equipment. Initial problems included duct replacement and supply delays, dust, and other personnel hazards. Operations analysis and SOP'S reduced delays during the second field trial, minimized hazards, doubled production. Re cent re search and operations analyses produced data on abrasion resistant ducts, identified minor equipment modifications which can improve production, developed techniques for impact compaction to 90% ?maximum? and for extension of the practical range to 2000 feet. Twenty-two illustrations include an air consumption-distance chart.

Jan 1, 1967

By R. Brantley Sudderth

There are three broad categories in which most solvent extraction technology can be distributed?: 1) reagents, 2) equipment, and 3) systems. Reagents which are now used in commercial solvent extraction plants are of many chemical types, and each possesses certain properties which make it more economically desirable than another alternative. The equipment which is' used in solvent extraction plants for contacting the organic and aqueous phases and separating the same is varied. Most common to the industry is the mixer-settler unit which in itself, can be of varied design; but many new innovations in contacting and separating equipment have been recently introduced or are :now being studied. Each specific solvent extraction application will often dictate the type of contacting and separating equipment used. Obviously, the trend is toward smaller less expensive equipment which reduces the often expensive organic inventories necessary for most solvent extraction plants. The current thrust of .investigations in ancillary equipment is usually aimed at recovering organic from solvent extraction sludges or from aqueous streams exiting the process.6 This type of equipment helps to reduce the total organic losses from solvent extraction plants and thus reduce the operating costs. The term "systems" in solvent extraction applies to the manner in which a solvent extraction process is carried out. This manner is usually dictated by the properties of the reagent which is used, and it involves a tailoring of the equipment or the process flowsheet. Some of the things which can be tailored are: 1) The aqueous media -whether acidic or basic and what type of acid or base 2) Counter-current or single stage (co-current) contacting 3) EMF or pH control 4) Organic "crowding" alternative 5) Heating or cooling process streams Regardless of the type or extent of tailoring, efforts to do this are usually a result of the low cost of solvent extraction as a unit process.

Jan 1, 1977

By P. C. Kapur

For mineral processing spirals, mathematical force equilibrium models are sought to extend the spiral performance envelop. In Part 1 of this report, a complete geometric description of the helical spiral and its trough or channel contour is provided. The secondary circulatory fluid motion in the transverse direction that develops when water flows down a tightly turning curvilinear path and its implications for gravity separation are discussed. Different forces that act on the particle during its passage through the spiral are enumerated and quantified for a better appreciation of their relative contributions and roles in the separation process. Three force equilibrium models from the hydrology literature on sed iment transport are employed to simulate a typical spiral. While the results are promising in that the models mimic the broad performance characteristics of the spiral, what is needed is a more general spiral model taking explicit cognizance of the missing components essential to the spiral description. These are such components as centrifugal force and the steep tangential slope of the channel deck. Results of this latter exercise are deferred to Part 2.

Jan 1, 1996

By M. S. Moats, W. C. Hartzell

The stable domestic supply of gallium, indium, and germanium has been identified as a critical need for the United States. Electric arc furnace dusts, processed to recover zinc, produce a residue that could be a potential source of these critical materials. An industrial zinc leach residue was characterized to determine its particle size, chemical composition, and predominant mineral phases. Heated, atmospheric, agitated leaching using sulfuric acid and hydrochloric acid was examined to determine the extractions of gallium, indium, and germanium as a function of concentration and temperature. Extractions of gallium were found to be similar to iron in all experiments. The maximum percentage of Ga, Ge, and In leached into solution were 71%, 63%, and 48%, respectively.

Feb 1, 2023

By Solomon J. Salat

During the major expansion of the Inspiration smelter, which took place in 1973, the bulk material handling system underwent a matching expansion in order to move the required materials. Inspiration is the first smelter worldwide to employ electric smelting in conjunction with the use of Hoboken converters (1), (2). The installation includes a 51 -MVA electric furnace, five 4.25m (14-foot) diameter Hoboken converters, rotary dryer, bedding plant, acid plant, heat recovery steam plant, and gas and dust handling system. See Figure 1, Plot Plan. As of today, the 51-MVA electric furnace is the world's largest in the smelting of copper. Engineering design for the installation was per- formed by Treadwell Corporation of New York. The task of designing the conveyor system was complex because the system had to be installed while the smelter was operating and with minimum plant shutdowns, and it had to serve both the existing and the new smelter. Existing conveyors were redesigned where necessary and incorporated into the new conveyor system. See Figure 2, Flow Sheet. The system encompasses the handling of such diverse materials as wet and dry copper concentrates, cement copper, silica flux, cold dope, and concentrate and flue dust.

Jan 1, 1979

By Carl E. Bagge

It is easy for a coal spokesman to relate to the theme of this meeting--that the decade of the 1970's is pivotal for the mineral industries. For the coal operator, this period is crucial. The coal industry, according to most energy forecasters, must gear its production machinery to meet a sharp rise in the demand for coal--in this decade to sustain inevitable boosts in electricity output, and in the next to start supporting a big new domestic synthetic fuels industry that is already in the formative stage. The logic of that projection flows from a combination of hard circumstances--proved reserves of domestic fluid fuels are dwindling, the hope for bountiful and cheap nuclear power must be indefinitely deferred, and the threat of long-term fuel imports to the national interests is becoming starkly visible. Only coal, with its assets of abundance, security and versatility, offers the nation an adequate and immediate lift out of its current energy crisis. A hard-won expansion of coal production made the critical difference between local power load-shedding and major area blackouts in 1970, the year the energy crisis went public. Coal will have to play that same key role, not just through this pivotal decade but through the turn of the century. But the logic of fuel demand apparently does not always extend to supply. The coal industry has had a huge production job laid out for it by the energy statisticians, but up to now the energy policy-makers have not followed through with the aids and inducements that the coal operator urgently needs to do that job well and completely.

Jan 1, 1972

By Manuel Gomez

This, report presents correlative data on ash and sulfur forms for Pittsburgh seam coal from Greene County, Pa. Prediction equations were developed for sulfur form distributions and washability characteristics. The equations developed were evaluated with an independent set of exploration core samples. It was demonstrated-that both sulfur distributions and washability characteristic's can be predicted from exploration core data ahead of mining. This information can improve the efficiency of selective mining techniques for the production of a uniform metallurgical coal product.

Jan 1, 1967

By J. P. Rider, T. W. Beck, J. A. Organiscak

"The National Institute for Occupational Safety and Health (NIOSH), Pittsburgh Mining Research Division (PMRD) has recently developed a series of models utilizing computational fluid dynamics (CFD) to study gas distribution around a continuous mining machine (CM) with various fan-powered flooded bed scrubber discharge configurations in an exhaust curtain working face. CFD models utilizing species transport model without reactions in FLUENT were constructed to evaluate the redirection of scrubber discharge toward the mining face rather than behind the return curtain. The study illustrates the gas distribution in the slab (second) cut. The following scenarios are considered in this study: 100% of the discharge redirected back towards the face on the off-curtain side of the continuous miner; 100% of the discharge redirected back towards the face, but divided equally to both sides of the machine; and 15% of the discharge redirected toward the face on the off-curtain side of the machine, with 85% directed into the return. These models were compared to a model with a conventional scrubber discharge where air is directed away from the face into the return. The models were validated based on experimental data and accurately predicted sulfur hexafluoride (SF6) gas levels at four gas monitoring locations. One additional prediction model was simulated to consider a different scrubber discharge angle for the 100% redirected, equally divided case. This paper describes the validation of the models based on experimental data and the gas distribution results of the CFD models. INTRODUCTION To control methane gas and dust concentrations on continuous mining operations, it is important to provide sufficient quantity of fresh air from the end of tubing or curtain to the active face. In this region, operating machine-mounted water sprays and fan-powered, flooded bed scrubbers can significantly increase the amount of airflow reaching the face [1-3]. However, when an entry is large, especially in higher coal seams, the water spray configuration and the scrubber capacity may not be adequate to maintain sufficient forward airflow to confine the dust cloud to the face. Dust can roll back toward the machine operator as a result [4]. Currently, several coal mines having negligible gas emissions in higher coal seams (2.4 to 3.0 m [8 to 10ft.] thick) with exhaust tubing systems redirect the scrubber discharge back towards the face in an effort to improve dust collection. NIOSH studied the redirected scrubber exhaust techniques for dust and gas control in a full scale CM gallery to assess its potential for industry-wide application [5]. For a CM face, previous studies have concluded that the behavior of gas and respirable dust is a complex process. The distribution/dispersion and transport of gas and respirable dust after they are liberated at the face are governed mainly by their mass properties, the movement pattern of the ventilation air, and interactions with the mining machine and water droplets created by the external water spray systems used for dust control [6-10]. To understand these behaviors in a complex CM environment and to evaluate the effectiveness of various control techniques, numerical modeling has become a necessary supplement to laboratory experiments and field studies."

Jan 1, 2017

By Donald R. Williamson

The Cyprus Pima Mining Company's open pit operation was started in 1957 with the normal assumption that the ore reserve estimate was accurate within an acceptable range of values. At that tine reserves were law - volume (+10 million tons) and high grade (+1.5% Cu). Several periods of growth have brought the operation to its present production level of 55,000 tons per day of ore and +125,000 tons of waste. Over the years, a number of attempts to refine and polish the ore reserve figures have been made using a variety of established standard procedures. None of these attempts have produced results which are totally satisfactory, particularly when comparisons are made between ore reserve estimates and short to medium range production figures. Other refinements were tried using adaptations of established methods combined with computer modeling techniques. The resulting mine-model ore blocks were an improvement but they still tended to show too match tonnage of mill feed, often at too low a grade. One attempt at Kriging had only limited success. Statistical studies of operations data revealed relationships which could be used to subdivide ore blocks in much the same way as they are actually mined. Comparisons thus far between these newly developed ore blocks and production data show that the new method of evaluation yields excellent results. A new, better procedure for estimation of ore reserves apparently has been developed for the Cyprus Pima deposits. The specific procedures seem likely to have direct application to other, similar operations. The methods used to develop the procedures could be of value to still other operations.

Jan 1, 1976

By Y. C. Kim

Using an exhaustive data set of Walker Lake, four different kriging techniques plus the inverse distance squared weighting were compared with regards to each one's performance in accurately estimating the block grades. The North-West quadrant of the Walker Lake data area was used because of its extreme variability of values and its distinct contour of the values. There were 1900 exhaustive data points in the selected area from which a total of 121 samples were drawn following the sampling strategy given in Isaaks and Srivastava (1989). The average of 5 x 5 exhaustive data points was treated as the true, known block grade. A total of 780 blocks, each block corresponding to the average of 5x5 exhaustive data points, were then estimated using the 121 sample points. The resulting estimates were then compared with the true known values to determine each technique's prediction performance. For this particular portion of the Walker Lake and for the given sampling strategies used, the obtained results for the global estimation are mixed. For the variable U which is much more variable, the rankings are; 1) Jackknife Kriging (JK), 2) Indicator Kriging (IK), 3) Inverse Distance Square weighting (IDS), 4) Ordinary Kriging (OK), and 5) Log-Jackknife Kriging (LJK). For the variable V, the results are; 1) OK, 2) IK, 3) JK, 4) LJK, and 5) IDS. When we apply several cutoff grades to the estimates to determine the conditional mean grades and tonnage, the above rankings generally get shifted around. Two exceptions to these cases are; a) UK for U and b) IDS for V. These two consistently ranked at the bottom.

Jan 1, 1995

By George A. Savanick

Borehole (slurry) mining, in which a tool incorporating a water-jet cutter and a downhole slurry pump mines phosphate through a borehole, was field tested in northeastern Florida as part of a Bureau of Mines program to increase the domestic availability of critical minerals. Approximately 1700 tons (1.54x106 kg) of phosphate ore were mined at 40 tph (10.1 kg/s) from a depth of 250 ft (76 m). The borehole was filled with water during mining for roof support and to assist slurry pumping. Underwater cutting was enhanced with an air-shrouded water jet. Mined cavities were backfilled to prevent subsidence after abandonment. Roof subsidence during mining was avoided only when the cavity was filled with water. Environmental monitoring before, during, and after mining included hydrological studies and measurements of water quality. Significant water quality changes occurred only in the zone being mined and were confined to the immediate area of the borehole.

Jan 1, 1983

By G. V. Sullivan

The Federal Bureau of Mines examined tailings samples from 13 western copper operations to determine the potential for recovering rutile. Results showed that six of the samples contained sufficient liberated titania to be considered potential sources. One sample contained 0.75 percent Ti02 or 0.40 percent liberated rutile. Batch flotation studies developed a procedure that recovered 70 percent of the liberated rutile at a grade of 34.7 percent TiO2. Sulfuric acid leaching studies on 12 of the samples dissolved 100 to 2,200 grams Cu per ton of tailings; leaching of the slimes showed slightly more copper was dissolved, 1,800 to 3,300 grams Cu per ton of minus lo-micrometer material. Bulk sulfide flotation of nine of the samples resulted in concentrates that contained 0.38 to 2.0 percent Cu and 0.7 to 2.1 and 6.9 to 42.9 grams per ton Au and Ag, respectively.

Jan 1, 1981

By Subhas G. Malghan

The response of feldspar flotation to tallow diamine dioleate collector in the presence of sulfuric acid is studied using batch flotation experiments. Application of this non-hydrofluoric acid (non-HF) reagent system in the flotation of mineralogically and chemically different feldspar ores is compared with that of the conventional hydrofluoric acid system. The dependence of flotation recovery and concentrate grade on feed particle size distribution, conditioning time, pulp density, and pH, combined variation of collector and acid addition rates, sodium hydroxide concentration in the collector solution, and initial pulp pH prior to conditioning is studied. The results have been discussed in relation to applicability of the non-HF system to plant trials, and the nature of the collector adsorption in the selective flotation of feldspar.

Jan 1, 1978

By Mahmoud Sepehrmanesh, Verya Nasri, Maziar Partovi

"This paper presents an assessment of the impact of tunneling and its associated risks on pile foundations located in the influence zone of tunneling operation. Tunneling in proximity of pile foundations introduces deformations in piles and in turn induces additional axial and bending forces. In modern mechanized tunneling, ground convergence is controlled through balancing the earth pressure at the tunnel face as well as applying pressurized bentonite through the shield. Reducing ground deformation can consequently lower the additional deformations and forces developed in the pile foundation and lessen the associated construction risk. In this paper, comprehensive detailed finite element modeling was performed to quantify the effect of the different parameters on induced forces in piles. Finite element models took into account the nonlinear behavior of soil and pile-soil interface. The modeling results are of practical importance in risk evaluation of tunneling in vicinity of pile foundations. INTRODUCTION Tunneling operation in congested urban areas becomes more frequent as demand for further utilizing underground spaces grows. In urban settings, it is not uncommon for tunnel designers to plan a tunnel route in close proximity to other structures to satisfy other competing objectives like raising the alignment to reduce excavation volume. Setting the alignment close to existing structures increases the risk involved in the project. It is the responsibility of the designer to thoroughly identify, assess, and mitigate all design risks involved. In this paper, we focus on the aspects involved in tunnel construction in proximity to existing pile foundations. Tunnel operations introduce deformation in the soil mass. Consequently, a pile located within deformation zone experiences extra forces as a result of the imposed deformation. The developed forces and bending moments can be in orders comparable with the pile capacity depending on the relative horizontal proximity of the tunnel and the relative depth of tunnel with respect to the pile length."

Jan 1, 2016

By R. F. Davis

The evaluation and selection of coals for producing metallurgical coke is indeed a com- plex process simply because coal is a complex and variable substance which is sensitive to how it is treated from the time it is mined until it is transformed into coke. A review of the the more common evaluation tests made on coal and how the test results are applied in commercial practice. Parameters used to select coal blends for wet and preheated charges should also be of interest.

Jan 1, 1980

By F. D. Ayhan

The aim of this study was to investigate desulfurization and de-ashing of a mixture of subbituminous coal and gangue minerals by the agglomeration method. For this purpose, experimental studies were conducted on a mixture containing subbituminous coal, pyrite, quartz and calcite. The effects of some parameters that markedly influence the effectiveness of selective oil agglomeration, such as solid concentration, pH, bridging liquid type and concentration, and depressant type and amount, were investigated. Agglomeration results showed that the usage of various depressants (Na2SiO3, FeCI3, corn starch, wheat starch) in the agglomeration medium has a positive effect on the reduction of ash and total sulfur content of agglomerates. It was found that an agglomerate product containing 3.03% total sulfur and 25.01% ash with a total sulfur reduction of 56.71% was obtained from a feed that contained 7% total sulfur and 43.58% ash when FeCI3 was used in the agglomeration medium. Key words: Desulfurization, De-ashing, Agglomeration, Coal, Gangue minerals

Jan 1, 2009

By M. L. Free, N. Dhawan, M. S. Safarzadeh, J. D. Miller, M. S. Moats

"The drum agglomeration of nickel laterite and gold ores has been optimized through the design of experiments (DOE) using a Taguchi L16 (45) orthogonal array to determine the optimum conditions for maximizing average agglomerate size and minimizing the amount of fines. The effects of controllable operating factors including moisture content (nickel laterite ore: 34-37%; gold ore: 7-10%), retention time (2-3.5 min), drum speed (15-45% critical speed), drum load (nickel laterite ore: 8-32 %; gold ore: 6-22%) and acid concentration (150-600 g/L) on the performance of the agglomeration process were studied.For nickel laterite ore, maximum average agglomerate size and minimum percent fines (-1 mm) occurred under the following conditions: drum load (23.7%), moisture (36.5%), time (3 min), drum speed (30% critical speed) and acid concentration (150 g/L). Under the studied nickel laterite ore conditions, the most effective parameters for maximizing average agglomerate size and minimizing the amount of fines were found to be drum load and acid concentration, respectively. Drum speed had a statistically significant effect on minimizing the amount of fines. Maximum average agglomerate size and minimum percent fines (-1 mm) for gold ore occurred under the following conditions: drum load (19.3%), moisture (8.5%), time (2 min 15 s) and drum speed (40% critical). The most significant factors for maximizing average agglomerate size and minimizing the amount of fines for gold ore were found to be drum load, time and moisture.IntroductionSeveral heap leach operations have experienced problems in terms of poor recovery due to percolation issues caused by low-grade complex ores, tailings and clayey deposits. Poor percolation can lead to low metal extraction due to solution channeling or the development of impermeable/dead zones within the heap (Dhawan et al., 2012; Kappes, 2005). Improper heap building practices such as the use of loose agglomerates and inadequate attention toward the presence of clay minerals have been among the main reasons for percolation issues. During the transport of ore material, severe segregation of the material can occur. To overcome percolation problems, a major improvement was made through the introduction of agglomeration prior to ore placement. If the ore particles and agglomerates are of similar size, segregation can be avoided to a great extent (Dhawan et al., 2013; Kinard and Schweizer, 1987; McClelland and van Zyl, 1988). Agglomeration improves the uniform percolation of solution through ore heaps and is applicable to many ores, wastes and milled tailings (Bouffard, 2005; Dhawan et al., 2013). Manning and Kappes (2005) reported agglomeration/ stacking accounts for ~14% of the total heap leaching operating cost. The cost of binder is a primary contributor toward the total cost. Also, the lack of consistent quality control tests for agglomerate often leads to operational problems. The details are mentioned elsewhere (Dhawan et al., 2013). Considering the significant contribution of the agglomeration/stacking step toward the total heap leach operating cost, it is appropriate to study the agglomeration behavior and its optimization for two different ores. The reason to study the same parameters for different ores is the fact that significant variability in the agglomeration behavior of different ores has been reported to exist in industrial operations."

Jan 1, 2015

By J. R. Dyni

Colorado nahcolite deposits, estimated at 29 billion metric tons, formed in an asp metric permanent-lake basin whose waters were rich in algae and sulfate-reducing bacteria. Nahcolite (NaHCOg) occurs in oil shale of the Green River Formation (Eocene) as aggregates (62 percent), disseminated crystals (24 percent), and impure beds (14 percent). Because much of the nahcolite is nonbedded, it will be mined with oil shale and can be concentrated to marketable grade by crushing and screening Impurities in nahcolite are kerogen, quartz, pyrite, and carbonate minerals. A promising new market for nahcolite is in the process of removal of sulfur dioxide from industrial stack gases. When heated, nahcolite reacts with sulfur dioxide to form sodium sulfate. Differences in thermally induced microporosity may explain why nahcolite is more reactive than trona, a possible competitor with nahcolite for SO2 removal. Limited production of nahcolite at one or possibly two localities is expected in the next several years, but large-scale production is probably at least 10 years away.

Jan 1, 1980

By D. C. Yang

A practical and economical flowsheet has been developed and demonstrated by integrating the packed column flotation with the GranuFlow process for fine coal recovery. Orimulsion, which was used as an additive, served a dual purpose: first in the flotation and second in the dewatering and granulation steps. The test results on the coal preparation effluent sample show that Orimulsion is flexible and compatible with fuel oil and various frothers used in the packed column. Clean coal with ash contents of 3.0, 3.7, 4.3 and 5.1% were produced from a head sample of 33% ash at combustible matter recoveries of 72.4, 81.1, 87.5, and 89.2%, respectively. Other advantages include high throughput, low moisture, improved handleability, and reduced dustiness of the dewatered coal, thus providing a very significant economic Incentive for the integrated process.

Jan 1, 1998

By W. R. Reed, J. P. Rider

"The National Institute for Occupational Safety and Health (NIOSH) Office of Mine Safety and Health Research (OMSHR) has recently developed a series of models utilizing computational fluid dynamics (CFD) to study airflows and respirable dust distribution associated with a medium-sized surface blasthole drill shroud with a dry dust collector system. CFD models were constructed in ANSYS FLUENT Version 15. Previously run experiments conducted in the NIOSH full-scale drill shroud laboratory were utilized to validate the models. The setup values in CFD models were calculated using experimental data obtained from the drill shroud laboratory and measurements of test material particle size. Subsequent simulation results were compared with the experimental data for test scenarios, including 0.14 m3/s (300 cfm) bailing airflow with 2:1, 3:1, and 4:1 dust collector-to-bailing airflow ratios and 0.24 m3/s (500 cfm) bailing airflow cases with 2:1, 3:1 and 4:1 dust collector-to-bailing airflow ratios. For the 2:1 and 3:1 ratios, results showed the calculated dust concentrations from the CFD models were within the 95% confidence intervals of the lab experimental dust concentrations. This paper describes the methodology used to develop the CFD models, to calculate the model dust input, and to validate the models based on experimental data. Problem regions were identified and revealed by the simulation of a dry dust collector system. The simulation results could be used for future development of dust control methods for a surface mine blasthole drill shroud. INTRODUCTION In surface mines, production drilling is an essential process in blasting which fractures the hard rock overburden for removal. Holes with a predetermined pattern, diameter, and depth are drilled into the overburden to introduce explosives. During the drilling process, considerable amounts of respirable dust are produced. Past research conducting area sampling at blasthole drills has documented time-weighted average respirable dust concentrations ranging from 8.68 to 95.15 mg/m3 [1] and 1.04 to 52.30 mg/m3 [2]. These high dust concentrations, which may contain silica, can lead to respirable dust over-exposures for the drill operator, drill helper, and other personnel in the vicinity. These over-exposures can lead to silicosis, an occupational lung disease that has no cure and is ultimately fatal. In order to prevent the development of silicosis to mine workers, the Mine Safety and Health Administration (MSHA) regulates the exposure to respirable dust. For metal/nonmetal (M/NM) mines, MSHA has adopted the threshold limit values recommended in 1973 by the American Conference of Governmental Industrial Hygienists (ACGIH) [3]. The total airborne dust standard is 10 mg/m3. However, if the sample contains greater than 1 percent quartz (the most common form of crystalline silica), a respirable dust standard is calculated with the following equation: Respirable dust standard = (10 mg/m3) / (% respirable quartz + 2)"

Jan 1, 2016