Search Documents

By J. E. House, C. J. Lewis

In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-pregnant sulfuric acid leach solution is contacted with an organic extractant dispersed in kerosene; the uranium enters the organic phase and associates itself with the organic extractants. Simultaneously an-ions or cations, as the case may be, leave the organic extractant phase and enter the aqueous phase. The two phases are then allowed to separate by virtue of their immiscibility. Based on the data available, a generalized procedure for the recovery of molybdenum from the sodium carbonate scrub liquor of acid leach uranium mills using secondary and tertiary amine extractants is presented, as is another generalized procedure for the recovery of molybdenum from the bleed stream of alkaline leach uranium mill circuits. Such economic appraisal as can be made at this stage of the development is also indicated. Molybdenum occurs in most uranium ores. Whether the uranium is solubilized by the acid leach route or by the alkaline leach process, some of the molybdenum present in the ore is solubilized along with the uranium. Molybdenum in uranium mill pregant solutions may cause processing difficulties, among which may be mentioned the poisoning of resins, the precipitation of organo-molybdenum complexes, or the build-up of molybdenum in recycle streams. While the uranium milling industry has been successful in applying measures to overcome the problems arising from the presence of molybdenum in the circuits, such measures nevertheless add to the cost of producing uranium yellow cake concentrate. Table I shows the analysis of typical uranium pregnant leach solutions. Uranium recovery by solvent extraction is presently practiced only in conjunction with acid leach circuits, and it may be noted that of about 16,000 tons of uranium ore daily processed by sulfuric acid leach, more than half of this tonnage follows the solvent extraction route. Uranium mills leaching ore with sodium carbonate-sodium bicarbonate solution (the so-called alkaline leach circuit) do not use solvent extraction. In this case the uranium pregnant solution is sufficiently pure to permit the direct precipitation of uranium concentrate with subsequent recycle of the alkaline leach solution. While solvent extraction procedures may be developed which are economically attractive for the alkaline leach uranium process, this has so far not been the case. The earlier uranium mill solvent extraction circuits used organo-phosphate extraction reagents, among which di-2-ethylhexyl phosphoric acid was popular. It should be borne in mind that uranium in its sulfuric acid leach solution exists in an equilibrium involving uranium in both anion and cation forms: uo + 2S0 = [uo2(so4)2] For this reason uranium may be extracted by either cation extractors or anion extractors and this is likewise true of molybdenum. However, the organo-phosphate extractants also load ferric iron which is usually present in the uranium acid leach solution. For this reason, it is necessary to reduce this solution substantially to the ferrous state ahead of solvent extraction using organo-phosphate reagents. The amines, being anion extractors, selectively reject iron and thus render the iron reduction step unnecessary. In the solvent extraction process, also referred to as liquid-liquid extraction, the clarified uranium-

Jan 1, 1961

By B Belle

In a predrained gassy coalmine, ventilation control plays a secondary role to the longwall (LW) goaf drainage systems to safely manage panel return gas levels. Typically, the LW specific gas emission (SGE) estimation models provide anticipated goaf gas emissions and arriving at required gas drainage capacity for various coal production scenarios. A study comparing the various LW goaf gas hole spacing patterns with weekly production rates of up to 270 kt/w of an operating mine with high SGE rates (~20 m3/t) was carried out. The goaf capture efficiency of LW panels ranged from 64 per cent to 85 per cent. For the first time, gas content of upper seam data post LW retreat were used to verify the current SGE prediction models. This study has shown that an average of 50 m goaf spacing is optimal for management of high production retreat gassy LW panels in presence of multiple upper and lower seams with gas emission zone influence reaching up to 200 m above workings. Due to the horizontal stress direction related geotechnical issues, goaf hole failures just behind the LW face is a recurring challenge resulting in significant and immediate increase in LW panel return gas levels. This paper highlights the shortcomings of current SGE prediction models for goaf gas estimation, and evolution of 50 m goaf hole spacing embedded in design process in managing the LW tailgate gas levels.CITATION:Belle, B, 2017. Optimal goaf hole spacing in high production gassy Australian longwall mines – operational experiences, in Proceedings Australian Mine Vent Conference 2017, pp 15–26 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 28, 2017

By J R. Lang, J K. Mortensen

The Canadian Cordillera contains numerous classic porphyry deposits and several unusual subtypes, including the alkaline suite of porphyry Cu-Au deposits and the Tombstone magmatic suite and associated Au¦W deposits. The alkaline suite formed in collisional settings in island arcs whereas the Tombstone suite was emplaced in an active fold and thrust belt in a continental arc. In addition to different regional settings, these suites differ in the type of underlying basement, metallogeny, and the styles of mineralization. Both magmatic suites were emplaced over short intervals, probably in response to specific tectonic events. In the case of the alkaline suite, the tectonic event promoted magmatic processes suitable for the formation of porphyry Cu-Au deposits. The tectonic environment is less well constrained for the Tombstone suite, but the result was the formation of distinctive porphyry and intrusion-related Au¦W deposits.

Jan 1, 1995

By Jon R. Maki

"EVTAC, an iron ore mine in Minnesota, had trouble meeting the customer’s pellet silica requirements. In addition, the mine plan was moving towards a lower grade ore and funding was not available for plant improvements to compensate for the change in ore or improving quality. Small changes to optimize various processing equipment resulted in significant improvements in both quality and productivity; for example, productivity on one mill line increased by 10% while producing lower silica concentrate. This paper will provide the methodology that allowed EVTAC to improve quality and productivity through process optimization with little capital investment.INTRODUCTION EVTAC, an iron ore mine in Minnesota, had trouble meeting the customer pellet silica requirements. In addition, the short-term mine plan was changing to a lower ore grade. Because of limited funds, EVTAC did not have the capital for investing in the plant to compensate for the ore change or to make improvements to lower product silica levels. Small changes to optimize process equipment resulted in significant improvements in quality and productivity; for example, one mill line feed rates increased by 10% while also producing a higher quality concentrate. These changes were sufficient enough to allow the mine to lower the target ore grade beyond their initial goal. At one time, a $3M flotation facility was considered as a way to meet quality with the new ore grade. However, the process changes were able to accomplish the same silica reduction with roughly a $50,000 investment in VFDs for sump controls. This paper will discuss some of the modifications that allowed EVTAC to meet the quality and production objectives with little capital investment. It also discusses the methodology used to make these process improvements."

Jan 1, 2014

By J. L. Howell, D. E. Pollock, A. B. Cecala, A. D. O’Brien

Dust from worker,s clothes has been shown to be a problematic source of personal dust exposure. A recently completed effort between NIOSH and the Unimin Corporation has resulted in a quick and effective way to clean worker clothing. The process involves a booth under negative pressure with an air spray manifold to supply compressed air to blow off the dust from the clothing. The overall system is designed to meet MSHA and OSHA requirements. Results of field testing indicated that the manifold cleaned the clothes 10 times faster and removed 50% more dust than cleaning methods used by workers today.

By Mohammed Sakr

A Two-Section Helical Pile (TSHP) is a steel pipe with two shaft sections of different diameters and with one or more helix plates near the bottom of each shaft section. TSHPs are used regularly in western Canada as foundation pockets for single and double circuit distribution power line poles. The results of a comprehensive lateral pile load test program and the field monitoring of TSHPs installed in organic soil (muskeg) over soft clay are presented in this paper. A total of nine full scale lateral load tests were carried out, including seven tests using TSHPs with different diameters and embedment depths, and two tests using one-section helical piles. The prime objective of the study was to evaluate the lateral resistance of TSHPs installed into very soft organic soil (muskeg) overlying soft clay soils, to compare their lateral resistance relative to one-section helical piles and to compare the measured and the estimated lateral resistances using p-y curves. In order to provide sufficient soil data for analysis, the subsurface soil stratification was evaluated using two different techniques including; Cone Penetration Tests (CPT) and Portable Soil Probe Tests (PSPT). The two testing techniques were also compared and evaluated. This paper summarizes subsurface soil conditions, pile configuration, helical pile installation procedure, testing setup and results.

Jan 1, 2009

By Hubert E. Risser

Although, for about a half-century the percentage of coal production provided through strip mining has steadily increased, recent trends indicate that a peak in percentage (not tonnage) is being approached, principally because of the decline in economic adrantage enjoyed by strip mining compared to underground mining. Factors contributing to the decline are the decrease in the margin of advantage in manpower requirements held by strip mining; an increase in depth and volume of overhurden, requiring larger equipment and capital investment; increasing lack of reserve land blocks large enough to support extensive strip mining operations; more stringent legislation for reclamation of stripped land, with increasingly higher costs of such reclamation; and the increasing costs of land surface rights for strip mining. Already there has been a peak and suhsequent decline in relative importance of strip mining insome major Eastern coal-producing states; trends indicate that a peaking for the whole Eastern region is on the way. Although, in the West the pressures are less severe, they will, nevertheless, reduce to an increasing degree the advantage that strip mining has possessed. For about half a century, strip mining has been providing an ever-increasing share of total coal output in the United States, and in the past 30 years this share has grown from less than 8% to 33.7%, as shown in Fig. 1. Fig. 1 also shows that strip mining is gaining faster in the West than in the East. From 1961 to 1967, output from strip mines grew 63.6 million toqs, or 52.2%. Certainly there is nothing in such performance to indicate any peaking in the level of strip coal-mine production. But the fact is that, if not in actual tonnage, strip mining appears to be reaching its peak in relative importance, at least east of the Mississippi River. And the area east of the Mississippi provides 90.0% of the strip and 94.9% of the total coal production of the nation. The relative economic advantage which spurred strip mining's growth for a third of a century has been diminishing markedly in recent years. Fig. 2 shows evidence of this decline. From 1947 until 1961, strip tonnage output remained fairly stable, at the end of the period showing only a 12.5% decline. Meanwhile, underground coal production fell 44.5%. As the result of the greater decline in underground coal, strip coal in 1961 furnished 30.3% of the total coal produced, compared to only 22.1% in 1947. In the succeeding six years, however, underground coal output increased by 76.5 million tons, while strip output increased only 63.6 million. Further evidence of the strengthening of underground coal's relative position — or if you will, the weakening of strip coal's position — can be found in the major coal stripping states: Illinois, West Kentucky, Ohio, Pennsylvania, Indiana, and West Virginia, which in 1966 provided 80% of the nation's strip coal. Fig. 3 shows what has been happening in three states lying in the Appalachian field and in the total output of these three states. In the figure, the column height represents the output of coal mined by all methods. The number at the base of each column is the percentage mined by stripping. In Pennsylvania and Ohio, the percentage of strip mining increased between 1956 and 1961, but by 1966 had declined from the 1961 level. In West Virginia, a decline occurred during the first five years but was followed by a rise. In January 1968, Coal Age published an article listing the locations and capacities of new large mines in the planning and construction stages that would be operating by 1968-1971. The plans for West Virginia and Pennsylvania indicated all the new mines would be underground. Except to the extent they might replace existing under-

Jan 1, 1970

By Paul Laframboise, Nigel MacDonald, Mike Hallewell, Robert Thorpe

"Galmoy Mine, owned by Arcon Mines Limited, has operated since 1997, and is one of three operating lead/zinc mines in Ireland. Changes in ore type starting in mid 2002 created challenges to metallurgical operations at Galmoy. This resulted in unacceptable lead and zinc flotation performance, which prompted initiation of a formal plant review and a subsequent stabilization programme. This involved a structured review of operational practices, development of a new Standard Operating Procedure (SOP) which in turn led to significant changes in process control equipment needs in conjunction with extension of zinc flotation cell residence time.This paper will describe the metallurgy of the Galmoy operation. Reasons for the poor metallurgy experienced in 2002 will be described, as well as the solutions identified to overcome these problems. By the second quarter of 2003, these solutions had been implemented and recoveries had risen to record levels.INTRODUCTIONGeneralGalmoy Concentrator is located in County Kilkenny (Refer to Figure 1 below) and was constructed during 1996 and commissioned in 1997.The concentrator at Galmoy had been designed initially to treat 650,000 tonnes per annum of primarily CW ore type. The main mineral of economic significance was sphalerite (Typically 10 to 11% Zn) with some minor galena (0.7%Pb) in a dolomite (Magnesium Carbonate) host rock.The only other competing sulphide mineral was pyrite (Circa 3 to 5% Fe in Mill Feed. This CW ore type predominated the first 6 years of production, but in August 2002, the ore mineralogy changed due to significantly higher levels of pyrite rich G ore body being blended in with CW ore."

Jan 1, 2005

By D. Brün, T. W. Nattkemper, T. Kuhn, T. Schoening

Underwater exploration for marine resources using camera-equipped devices (OFOS, ROV, AUV) is receiving growing attention. This development is creating a continuously growing pile of digital underwater images leading to a serious bottleneck in the analysis. However, any harvesting strategy must be coupled to an efficient and effective assessment of the biodiversity in the area of interest and in nearby control areas. To integrate manual image and computational image interpretation on different levels (i.e. regarding geology and environmental sciences and taxonomy), sophisticated software support is required, which is not yet offered explicitly, neither by commercial software companies nor by academic or other publicly founded institutions. Here, we present the first version of a new digital asset management system that offers functionalities for easy data management, sharing, manual annotation and integration of computational modules for image enhancement/pre-processing and interpretation. The functionality is demonstrated with data from the CCZ provided by the BGR Hannover.

Sep 1, 2014

By William S. Sanner

A demonstrated reserve base (DRB) of bituminous coal suitable for cokemaking has been estimated based on the latest published information on U.S. coal reserves. Bureau of Mines analyses of raw coalbed samples were used to classify the coals as low-, medium-, and high-volatile bituminous and grade them as premium, marginal, and latent coking coals. Coal rank was used as the basis for identifying coking coals, and ash and sulfur content were used to determine coking coal grade. The demonstrated reserve base of coking coal for the entire United States is 162.67 billion tons. However, only 18.96 billion tons are premium (metallurgical) grade, and between 80 and 99 percent occurs in the Eastern States. The demonstrated reserve base of marginal-grade coking coal is 31.3 billion tons; more than half the marginal reserve base is in the Eastern States, and most of the medium- and low-volatile marginal coals are in the Eastern States. The Eastern and Midwestern States contain most of the latent-grade DRB of 112.4 billion tons; most (106.6 billion tons) is high-volatile coal, and 98 percent of the medium- and low-volatile latent coals occur in the Eastern States. Based on the worst assumption, which assumes that only premium-grade coals are used for cokemaking, the DRB for high-volatile coal would suffice for 127 years, the medium-volatile coal for 80 years, and the low-volatile coal for 83 years.

Jan 1, 1979

By N. Abdel-Khalek, A. Youssef

The tailings from beneficiation of Florida phosphate ore contain considerable phosphate values. These tailings are currently discharged into large disposal areas. More than three million tons of P2O5 are discarded with the phosphatic slimes (clays) in Florida each year. These phosphatic clays not only create one of the most difficult disposal problems in the mining industry, but also represent a tremendous loss of P2O5. As the quality and quantity of Florida’s phosphate reserves diminish, recovery of phosphate values from these wastes becomes more vital. This will also significantly lengthen the life of these important phosphate resources. Thus, it is the goal of this study is to recover more than 80% of the phosphate values in phosphatic clays. A systematic study to characterize samples taken from inactive disposal area has been conducted. The data suggest that natural segregation takes place in the settling pond. In other words, coarser particles settle closer to the feeding point of the pond. Flotation tests were conducted using the +325 mesh fraction of samples obtained from sites close to the feeding point (sample# A) and further from that point (sample# B) and (sample# C) which is even further away from the feeding point. The results show the following: Sample# A gives the highest grade and recovery, followed by that of # B and then of # C. There is a great potential to upgrade sample# A where concentrates of high grade (68.9-70.1 % BPL or 31.5- 32.1% P2O5) and recovery (~86.6-95.4 %) can be obtained under the investigated operating conditions (1.0 – 2.0 kg/ton sodium silicate and 1.5—3.0 kg/ton collector mixture of fatty acid and fuel oil). The mineralogical and chemical analysis of all samples is presented in this paper for comparison. Only in this paper, the flotation data of (site # A) are discussed below.

Jan 1, 2002

By F. Baerhold, S. Mitterecker

"Acid recovery is an important process step in pickling and leaching processes. Pyrohydrolysis in spray roasters and fluidized beds is a well-established technology to recover hydrochloric acid, but the hydrolysis in the liquid phase is more and more considered as an alternative new process route. The characteristic features of these three processes will be discussed along with select industry trends. For example, hydrochloric acid based leaching processes for ilmenite and nickel laterite ores produce far larger volumes of waste acid streams than the steel industry. Scaling-up of the spray roasting technology to the highest possible capacity is necessary in order to become more cost effective. In addition, process routes based on ferric chloride pyrohydrolysis allow the production of super-azeotropic HCl (up to 30 wt.%). Finally, the new ECOmode concept optimizes the operation of both pickling lines and the spray roast acid recovery plant. Long term results from industrial applications show that fuel savings in the spray roaster of 25% are easily achieved without additional process steps. An attractive feature is also to upgrade existing installations to higher plant capacities.INTRODUCTION Pyrohydrolysis is the conversion of metal halides into metal oxides at elevated temperatures in the presence of water vapor. Reaction products are a hydrogen halide gas and a metal oxide. The process was originally developed and introduced into the market by ANDRITZ Ruthner in the 1960’s. It has found wide technical application in the field of acid recovery and oxide production. The introduction of this process has revolutionized pickling technology. The use of hydrochloric acid has strongly increased over the last decades, since pyrohydrolysis is available for the total recovery of waste pickle liquors from mild steel pickling, as illustrated in Figure 1.Chloride media are also used increasingly in the metallurgical industry. Oxide production processes for Mg, Co, Ni, and Al have been installed. Processes using hydrochloric acid are known for the leaching of ores as well as for the re-extraction of metals from organic solvents. In both cases the recycling of HCl is of substantial importance to achieve highly economic processes and to reduce waste."

Jan 1, 2016

By D. M. Weinberg, S. R. Novascone, R. R. Reynolds, J. H. Cole, C. D. Christensen

The mechanical properties of the rocks surrounding an orbital vibrator alter the shape and amplitude of the vibrator’s orbit. Detecting these changes with driving point impedance measurements promises to reveal fractures and, possibly, identify the in situ stress state and significantly increase the depth that can be investigated. Our research emphasis has been on detecting vertical fractures through casing. Experimental and numerical results to date indicate that our new logging tool has excellent potential for determining both the vertical extent and azimuthal orientation of such fractures. Fracture and in situ stress orientations are useful in mine planning.

Jan 1, 2002

By Budiselic C, Hefter G

The uptake of gold by activated carbon from acid chloride solutions involves the reduction of gold(IIl) to metallic gold. Deposition occurs only on the surface of the carbon at localized active sites at first, and subsequently at less active areas. The rate of uptake depends on the surface area of activated carbon, which confirms results from a recently published study (de Siegel and Soto, 1984). Increased rates of uptake were observed at higher temperatures and lower chloride concentrations due to the decreased stability of AuCI4-, the major gold(IIl) species present. Recoveries of gold approached 100 per cent for a contact time of 60 minutes at 20°C in a well stirred solution initially 5 ppm in gold(III) and containing 0.5 per cent w/v coconut-shell carbon. The presence of 0.25 Mnitric acid or 100 ppm iron(IIl) had no detectable effect on either rate or efficiency of gold uptake.

Jan 1, 1985

By JOHN W., E. C. SARCENT

The South Texas area discussed herein represents districts 2 and 4 of the Texas Railroad Commission. It extends from Jackson, Lavaca and Gonzales Counties on the northeast to the Rio Grande River, and comprises a total of 26 counties.

Jan 1, 1946

By C. K. Kocsis, S. Hardcastle

Mine ventilation, requires electricity to operate the fans and fossil fuel for heating. Typically, this accounts for approximately 40% of the overall underground energy consumption. This is considerable, and one of the major causes of this expense is that most mine ventilation systems are designed for the "worst-case-scenario" with respect to ventilation demand, which is commonly dictated by diesel equipment usage. Once designed, mines tend to operate their ventilation system at this peak level despite their exceeding day-to-day development and production needs due to lack of appropriate ventilation controls. In underground metal mines, TeleminingTM, which includes the tele-remote operation of mobile development and production equipment from surface, will change the mine design criteria including its support infrastructure such as the ventilation system. With the removal of machine operators to surface control rooms, conventional ventilation systems designed to dilute and remove pollutants can become redundant. Depending on the ultimate need for ventilation, its cost in automated mines with tele-remote equipment operated from surface will be lower than the cost of ventilation in conventional mines. These cost savings, which could be significant, would not only come from the ability to reduce the amount of ventilation in the mine, but also through the capacity to automate its distribution. This paper provides an indication of the new airflow requirements in stopes and dead-end developments for the tele-remote and automated mining processes based on climatic modelling. Furthermore, this paper makes a cost benefit analysis by comparing the ventilation system operating cost of a conventional underground metal mine based on the mobile diesel powered equipment design criteria with the newly redesigned ventilation system operating cost of an automated mine based on the "heat management" design criteria.

Jan 1, 2002

By Zephaniah Varley, Dominic Parmantier, Paul Schmall, Richard Martin, Red Robinson

Sequential excavation of the Beacon Hill Station has been accomplished in highly complex and inter-layered sand, silt, and clay soils with multiple perched groundwater tables. Successful excavation has required the installation of a deep dewatering well system used to depressurize sand layers above and within the excavated tunnel faces. In addition, the presence of sand layers in the crown of the Platform Tunnels required jet grouting at depths of up to 49m (160-ft). This paper discusses the design intent, construction, and effectiveness of the dewatering system and the selective jet grouting of sand horizons and shaft break outs for the successful construction of the deepest SEM station excavation ever completed in the United States.

Jan 1, 2007

By K. C. Armstrong, A. B. Cron, L. A. Melis

"The Johnny Mountain gold mine in northern British Columbia, owned by skyline Gold Corporation, has been in operation since August 1988. Mill operating difficulties at the start of the operation resulted in a depleted cash flow situation and quick improvements were required to generate a self-sustaining financial basis of operation. In October 1988, Melis Engineering Ltd. was contracted to identify metallurgical difficulties, implement necessary modifications, and undertake mill operations and personnel training in order to achieve production targets.This paper presents the operating difficulties which existed at start-up and describes the methodology used to overcome production problems. Metallurgical testing results are described along with the modifications made to the milling circuit to suit ore metallurgy. Plant changes led to the operation of a simple gravity/flotation circuit to recover gold as bullion and in a copper concentrate. Operation of the originally installed cyanidation circuit was discontinued. This contributed to substantial savings in operating costs and virtually eliminated environmental problems. PROPERRTY LOCATION ANR DESCRIPTIONThe Johnny Mountain mine site is located in northern British Columbia along the Iskuit River about 60 km east of Wrangell, Alaska and 275 km north of Terrace, B. C.. A location map is presented in Figure 1 and a site plan is shown in Figure 2. The complex is located on a mountain plateau about 3, 600 feet above sea level. The plateau is flanked to the north by Johnny Mountain and surrounded on the other sides by steep cliffs."

Jan 1, 1990

By Vance G. Leak

A technique for recovering copper and steel from metallic scrap has been developed by the Bureau of Mines based on preferential melting in molten salt baths at 1,150° to 1,250° C, The molten salt, preferably barium chloride or calcium chloride, both minimizes oxidation of the metals and promotes heat transfer. Cleanest separations (0.01 percent copper remaining in iron) have been achieved by pretreatment (dipping or spraying) with sodium sulphate or sodium silicate, which alters and coats the surface of the iron to inhibit the alloying or brazing with copper. An economic evaluation of the process was carried out on six combinations of sweating media and inhibiting reagents. On the basis of the postulated model, a cost of about $0.20 per pound of copper recovered has been determined for the sweating treatment operation.

Jan 1, 1973

By A. W. Khair

This paper deals with the design characteristics of cutting tools and the operating parameters these tools are usually subjected to. Eight types of cutting tools/bits with different geometrical shapes and sizes were tested in a linear cutting apparatus, utilizing concrete blocks. The objectives of these experiments were to study the performance of each type of bit, make qualitative assessment of each bit geometry and size and establish cutting parameters for efficient utilization of the tools. During the experiment required forces Fx, Fy, and Fz, cutting, penetrating and lateral forces respectively and audible and sub-audible noises were measured. In these experiments, the depth of cut was kept constant to 1.5 inch while the bit spacing was varied 1.5,2.25 and 3.0 inch (3.75,5.6 and 7.5cm respectively). The results of these experiments indicated that by optimum ratio of depth of cut to bit spacing, the cutting face will be cleaned with the least ridge/land remains between the bit groves, resulting in least specific energy consumption, more large and less fine product particles and substantial reduction of audible noise. The results of these experiments also showed that proper bit design not only decreased cutting energy consumption but also reduced audible noise.

Jan 1, 2001