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By J. M. Bray, R. R. H. Cohen, P. Kiousis, S. R. Al-Abed

"INTRODUCTION Conventional tailing storage and management practices typically utilize engineered facilities, such as ring dikes and valley impoundments. To reduce the environmental costs associated with disposing of mine wastes in tailing storage facilities (TSFs), several beneficial re-use strategies have been developed over the years. For example, thickened tailings have been used for paste backfill. It is common for tails to comprise up to 60% of the material’s total weight and it has been estimated that TSF volumes can be reduced by 60% when using this strategy, (Benzaazoua, et al. 2004). Other beneficial re-use strategies have produced unpaved temporary access roads, hot and cold asphalt, chip seals, self-compacting concrete, and golf course sand traps. This project sampled mine tailings from two different TSFs in Colorado and produced structural concrete that could potentially be used for commercial foundations, sidewalks, and paved roads. The tailings were substituted for the fine aggregates typically found in conventional concrete. The two main objectives of this study were to investigate the structural integrity of the produced concrete samples and the leaching of heavy metals from the resulting concrete. The ratio of tailings to fine aggregate (T/FA) was carefully studied to find if there was a correlation between the substitution rate and the structural integrity. The potential maximum substitution rate was also investigated for the produced concrete specimens. The potential of leaching mining impacted water (MIW) was evaluated by comparing the leachability of the raw mine tailings to the encapsulated tailings by using standard shake extraction prediction tests. MATERIAL AND METHODS Materials The four components that make up the concrete mix design for this project are: cement, coarse aggregate (gravel), fine aggregate (sand and mine tailings), and water. The type I/II ordinary Portland cement (OPC), gravel, and sand came from local cement and concrete companies in Denver, Colorado. Mine tailings were collected from two different TSFs in the Rocky Mountain Region, where Pride of the West TSF in Silverton, CO will be denoted as Site A. The primary metals of interest at the mill site are gold, silver, lead, copper, and zinc. The location of Site B will not be mentioned in this report due to confidentiality agreements with the mining company"

Jan 1, 2016

By Donald J. Reining

The California Surface Mining and Reclamation Act (SMARA) is the nation's first legislation establishing a statewide mineral resource policy governing mining activities that recognizes regional needs and retains local autonomy. The Southern California Rock Products Association supported this proposed legislation. It was through our efforts that SMARA was passed in 1975. The local planning and zoning processes, which are so critical to the mining industry, have been significantly improved with implementation of SMARA. Recognition of the importance of the mining industry is now being developed by the State Mining and Geology Board and the California Division of Mines and Geology (CDMG).

Jan 1, 1984

By S. Klein, M. McLeod, K. Johnson

The East Bay Municipal Utility District (EBMUD) delivers water to the San Francisco Bay Area through an aqueduct system extending from the Sierra Nevada Mountains to the East Bay. This aqueduct system, the Mokelumne Aqueducts, are comprised of three 82-mile long steel pipelines. The portion of the aqueducts crossing the Sacramento-San Joaquin River Delta is recognized to be vulnerable to seismic and flooding hazards. Previous studies evaluated alternatives for mitigating these hazards and increasing the reliability of the aqueducts for the next 100 years. The preferred solution involves construction of a 16.5-mile tunnel across the Delta and re-routing the existing aqueducts through this tunnel

Dec 1, 2022

By James E. Lawver

Response surface methodology is a well-known and powerful technique for determining optimum conditions in flotation systems. One disadvantage is the onerous task of the numerical calculations and curve platting. This paper describes a system in which the investigator simply selects the range of several flotation variables after which the design of the experiment, data analysis, generation of response surface equations, and response surface plats are automatically carried out using a digital computer and a computer-driven curve plotter. Several typical response surfaces of important variables in phosphate flotation are presented as examples. An overview of the flow charting of the system is provided in the addendum.

Jan 1, 1978

By A. Robison, V. T. McLemore

"The San Juan Basin, New Mexico contains many heavy mineral beach-placer sandstone deposits of Cretaceous age. Beach-placer sandstone deposits are concentrations of heavy minerals that formed by mechanical concentrations (i.e. settling) of heavy minerals on beaches or in longshore bars in a marginal-marine environment. They have high concentrations of several elements, such as Ti, Zr, rare earth elements (REE), Sc, Y, U, Th, Nb, Ta and Fe. These elements are becoming increasingly more important to industry, because they are used in the manufacture of devices our society uses every day, i.e. cell phones, computers, and alternative energy devices, like solar panels. It is unlikely that any of the beach-placer sandstone deposits in the San Juan Basin will be mined in the near future because of small tonnage, low grades, high iron content, and distance to processing plants and markets. However, as the demand for some of these elements increases because of increased demand and short supplies, the dollar value per ton of ore rises, enhancing deposit economics. Current studies at the New Mexico Bureau of Geology and Mineral Resources are focused on the Apache Mesa (formerly Stinking Lake) deposit near Dulce, New Mexico. INTRODUCTION Beach-placer sandstone deposits are accumulations of heavy, resistant minerals (i.e. high specific gravity) that form on upper regions of beaches or in long-shore bars in a marginal-marine environment. They form by mechanical concentration (i.e. settling) of heavy minerals by the action of waves, currents, and winds (Fig. 1; Bryan et al., 2007). Modern examples are found along the Atlantic Coast, USA (Koch, 1986; Carpenter and Carpenter, 1991), southeastern Australia (Roy, 1999), and Andhra Pradesh, India (Rao et al., 2008), where they are mined for titanium, zircon, and monazite (a Ce-bearing rare earth, REE, mineral). Detrital heavy minerals comprise approximately 50-60% of these sandstones and typically consist of titanite, zircon, magnetite, ilmenite, monazite, apatite, rutile, xenotime, garnet, and allanite, among other minerals. Most of these minerals have a high specific gravity exceeding 4 and are dark colored, giving the sandstones a dark color, resulting in them being called black sandstones. Although beach-placer sandstone deposits are found in strata of all ages; the deposits in the San Juan Basin in New Mexico are restricted to Late Cretaceous rocks belonging to the Gallup, Dalton, Point Lookout, and Pictured Cliffs Sandstones (Fig. 2; Table 1; Chenoweth, 1957; Houston and Murphy, 1970, 1977). The beach-placer sandstones are black, dark gray, to olive-brown, resistant to erosion, and radioactive due to radioactive zircon, monazite, apatite, and thorium minerals. Anomalously high concentrations of Ti, Fe, Nb, Th, U, Zr, Sc, Y, and REE are characteristic of these deposits. Similar Upper Cretaceous heavy mineral, beach-placer sandstone deposits are found throughout Montana, Wyoming, Utah, Arizona, and Colorado (Dow and Batty, 1961; Houston and Murphy, 1970, 1977; Zech et al., 1994)."

Jan 1, 2016

By Brian W. Mackenzie

The mine development decision is made on completion of exploration to determine whether or not a mineral deposit should be developed to production and, if so, how. This involves the selection of an optimum mine development alternative from a number of technically feasible alternatives and the comparison of this optimum with those of other available investment opportunities. If the optimum is sufficiently attractive, the deposit is developed to production. The selection process is carried out within a framework of corporate objectives and resources. The mine development decision is based on the balancing of two parameters - - rate of production as reflected in plant capacity and cut-off grade. On the basis of available information and experience, detailed revenue and cost 'estimates are made for each feasible combination of plant capacity and cut-off grade. What economic evaluation techniques should be applied to these estimates to reduce each alternative to economic indicators that will enable meaningful comparisons and an informed investment decision? This is the essence of the paper. The determination of optimum plant capacity and cut-off grade has been the subject of intensive research efforts by economists and mining engineers (1-7). Their studies are based on marginal analysis and assume that, at the time of decision, certainty conditions prevail with respect to ore reserves, mining costs and mineral markets. The optimum mine development alternative is defined as that which maximizes profitability.

Jan 1, 1969

Geostatistical literature has been dealing with recoverable reserve estimation techniques for the last 30 years. There have been well established support correction techniques (Le. affine correction, indirect log-normal correction, etc.) to deal with volume -variance relationship globally in determining what will be the reserves above a cutoff grade during mining. Kriging has become the accepted technique in the mining industry for ore reserve estimation. Many large companies are using this method for gold and copper deposit modeling. In general, the estimates coming from kriging are smooth. The variances of the block grades estimated by kriging are often lower than the material that will be mined selectively. As a result, kriging tends to under estimate reserves at high cutoff grades and overestimates them in low cutoff grades. It is necessary to perform additional steps after kriging if the goal is to determine recoverable reserves at some cutoff and not the in situ resources. In the mining industry, it is becoming more and more common for the low grade heap leach projects to consider whether the higher grade portion of a deposit may be processed by a mill at higher profits. Similarly, a concentrator -refinery operations are implementing heap leach -SX-EW option for the lower grade portion of their deposit. These options are making it very critical to estimate the low and high end of the grade distributions more accurately. Ore reserve estimation is normally done by using widely spaced exploration data which has very small size support (Le. core samples). Selective mining is carried out on ore control that is based on closely spaced blasthole data normally averaged into larger polygons based on some cutoff grade. Estimated grades from exploration data need to reflect the variability of the grades of the selectively mined units if they are used without further adjustments to predict recoverable reserves.

Jan 1, 1997

By R. V. Ramani, C. J. Bise

INTRODUCTION Classification of coal seams as thin, medium, and thick is both subjective and relative and is more a reflection of the statistical mode of the thicknesses en- countered in a mining region. A geological factor more precisely understood as defined than thickness is the inclination of the seams to the horizontal, commonly referred to as the pitch or the dip of the seam. Thick and steeply pitching seams present far more challenges to the ingenuity of the mining men to devise methods for their extraction than the thick and less steeply inclined ones. In a marginal case, where a thick coal seam may be minable by surface or underground methods, the decision bears a close relationship to the technological developments in the mining scene. Near surface thick coal seams are a special gift of nature and are not as wide- spread as the ones that call for the application of under- ground methods. Recent advancements in surface mining equipment have made it possible to surface mine to great depths. Stripping ratios of 30: 1 (cu yd of over- burden removed per ton of coal) have been achieved. The extraction of thick seams which preclude the use of surface methods is the concern of this chapter. Specifically, the objective here is to outline the special technological factors in underground mining of thick coal seams and to illustrate through several case histories the methods that are practiced in the North American continent. For the purposes of this chapter, a thick seam is defined as one which is not readily minable with currently available equipment without modifications to the extraction method or the equipment. At the present time, a thickness of 4.8 m (16 ft) can be considered as the lower limit, although developments in equipment and methods can be expected to push this limit upwards. The data for the case histories were collected from two extended mine visits in 1975 and 1978, and from other published accounts of the mines visited. Reference is drawn particularly to reports by Bise, Ramani, and Stefanko (1976) and Rarnani and Kenzy (1979). THICK SEAM MINING METHODS The various methods of mining thick seams can generally be grouped into three major categories: full face methods, where the entire thickness of the seam is re- covered in a single pass; slicing methods, where the seam is extracted to its full thickness by horizontal or inclined slices either in an ascending or descending order; and caving methods, where the seam is undercut and the en- tire coal seam is recovered by caving the top coal into the excavated cut, or some variations of the basic sub- level caving method. As shown in Table 1, in each of these methods, the coal is removed by either room-and-pillar or longwall methods. Hydraulic mining and transport of the coal in open flumes or in pipelines are also practiced. Addition- ally, the void created by the removal of coal may be filled with a packing material (stowing) except in the caving methods. The pitch (dip) of the seam, particularly when it is greater than 0.26 rad (150), imposes severe operational limitations on mobile equipment. A more detailed general discussion of the methods can be found in Bise et al. (1976) and Vorobjev and Deshmukh (1964). Specific discussion on the many variations of the three basic methods can be found in the papers presented at the international symposia that were held in

Jan 1, 1982

By Irina Melekestseva, Beltenev. Victor, Victor Ivanov

"One of the major problems discussed over the last years is a possibility of mining of seafloor massive sulfide deposits. In spite of the pessimistic points of view on their commercial development, they are still of great interest and continue to be studied (Hannington et al., 2011). Recently, Barriga (2012) paid special attention to the underestimated amount of massive sulfide hydrothermal fields and pointed to the importance of the buried fields and subseafloor mineralization.This study examines the geochemistry of seafloor and subseafloor precipitates collected at the Peterburgskoe (19°52´ N) and Irinovskoe (13°20´ N) hydrothermal fields (MAR) during the 34th cruise of the R/V Professor Logatchev in 2011. These fields were discovered by the Polar Marine Geosurvey Expedition in 2010 and 2011, respectively (Shilov et al., 2011; Beltenev et al., 2012). The Peterburgskoe field, associated with basalts, is located 16 km west of the main rift valley at the depth of 2800-3100 m at the junction of a large non-transform discontinuity with the second rift ridge parallel to the rift valley. The Irinovskoe field occurs at depths of 2700-2850 m and has a northwestern satellite sulfide site, both confined to an oceanic core complex (MacLeod et al., 2009), which includes mafic and ultramafic rocks. The details of geological structure of the fields may be found in the relevant papers.Diverse samples were collected from the hydrothermal fields: seafloor massive sulfides, Fe-hydroxide crusts, SiO2-rich sulfide-bearing precipitates, and subseafloor mineralized rocks (Melekestseva et al., 2014a). Massive sulfides may be subdivided on the Fe-rich and Cu-rich types. The marcasite-pyrite, pyrite, hydroxide-marcasite, and chalcopyrite-pyrite (Fe-rich) mineral types are dominant in the Peterburgskoe field at lesser amounts of sphalerite-hydroxide-chalcopyrite and pyrite-covellite-chalcopyrite (Cu-rich) massive sulfides. The mineralized sulfide-talc-carbonate-chlorite, chlorite-sulfide-quartz, and hematite-sulfide-quartz rocks host stringer-disseminated sulfides including chalcopyrite, covellite, and pyrite."

Sep 1, 2014

By Peter Cammack

Where raw coal from a mine is treated in dense medium separators ash in the washed product is generally very consistent, provided control of the medium suspension is maintained at a given specific gravity. When, however, the washing process is a Baum jig or other similar apparatus, making a separation at rather arbitrary gravities, the control of the product ash is less accurate, particularly in those cases where the design capacity of the washbox is exceeded by the load of high ash fines. In either of these cases an instantaneous measurement of ash while not essential would be desirable, to check the quality of the washed product and provide operator or management with instantaneous warning of changes in ash. Nearly 60% of all British coal mined goes for electricity generation, to burn at 17% ash, and since the overall average ash content of raw coal as mined is about 30% and varies within a range of 18 ? 40%, incombustible material has to be removed. The ash specification could be met either by total washing, or by washing part and blending the clean coal produced with the remaining untreated coal. Some plants use a straight tonnage split, but generally speaking in Britain it has been found more economical to screen and wash the oversize only. In either case for effective quality control in blending an instantaneous determination of the ash content of the product as it is made has been found essential, since standard ash determination by pyrolysis methods is lengthy and provides information too late to be used in the control of plants blending at the rate of 1000 ton/h.

Jan 1, 1976

By Amila Sandaruwan Ratnayake, T. Dilmi. U. Wijewardhana, H. C. S. Subasinghe

The value addition potential of ilmenite was examined using several characterization techniques. Raw ilmenite is composed of 93.65 wt.% of FeOTiO2, 3.55 wt.% of SiO2, 1.22 wt.% of Al2O3, and the remainder being other minor oxides. Ilmenite and powdered carbonized coconut shells were mixed in the weight ratio of 4:1. A separate fraction of the same mixture was added with powdered seashells in the weight ratio of 4:1:0.5. Six fractions of each mixture were distinctly milled from 1 to 6 h in a planetary ball mill. XRD spectra (broadened and diminished reflections of ilmenite) and FTIR observations (Ti―O―Cbonding) of milled samples indicate the possible incorporation of carbon into the ilmenite structure. Any clues of an occurrence of new rutile peaks were not observed in XRD spectra of milled samples. Consequently, the carbothermic reduction has not been initiated during mechanical activation. In this case, samples powdered for 6 h in a mill were isothermally annealed for 2 h under normal airflow at temperatures of 800 °C, 1000 °C, and 1200 °C, respectively. According to the X-ray diffractograms, the annealing temperatures of 1000 °C and 1200 °C exhibited almost similar trends with rutile (R), pseudobrookite (PB) and elemental iron (F), and a very few ilmenite (I) peaks. Consequently, 6 h of milling and 1000 °C annealing were concluded as the optimum conditions for the carbothermic reduction. Moreover, this study indicated seashells as a potential rate raiser for the carbothermic reduction of ilmenite at 800 °C. Therefore, this process is applicable to upgrade ilmenite into a mixture of synthetic rutile and elemental iron.

Mar 29, 2021

By F. A. Edwards

The Macassa Division of Lac Minerals Ltd. is a high grade gold mine that has been operating in Kirkland Lake, Ontario for the past 51 years. It produces approximately 120,000 tonnes of ore per year, averaging approximately 18.7 grams per tonne (0.45 ounces per ton) of gold. Over the years, the ore body has been developed to a point where mining takes place below the 1,220 m (4,000 feet) elevation and 1,830 m elevation (6,000 feet) west of the original shaft. The mine is currently serviced by two surface and two winze shafts. (Figure 1). The ore body extends below the 1,930 m (6,400 feet) Level which is the bottom of the existing mining operation. An ambitious underground exploration program has substantially increased these reserves. Concurrent with the discovery of greater reserves, has been rapidly escalating mining costs, particularly labour which constituted 66% of the total cost in 1981. These costs continued to escalate as the major proportion of mining was carried out to the west, far away from the existing hoisting systems. This required a three lift hoisting system for men, materials and ore. In addition to the increased travel time and ore handling expense, increased humidity, temperature, ventilation and compressed air losses have added to the inefficiency of the operation. As a result, studies were initiated in 1980 to assess the future of the Macassa operation. The results of the studies indicated that a new deep shaft with a potential to hoist from 2,135 m (7,000 feet) or below was the most attractive option. Detailed design of the shaft and associated surface plant commenced in January, 1982 and the following milestone dates were achieved: 1980 - initial study of hoisting and haulage facilities; January, 1981 - ore reserves in excess of 1 million tonnes of ore at 18.7 g/tonne (0.45 oz/ton) confirmed; March, 1982 - preliminary study of Macassa No. 3 Shaft completed; April, 1982 - permanent hoisting plant purchased; June, 1982 - site clearing commenced; September, 1982 - shaft collared; May 1, 1983 - shaft sinking commenced; February 11, 1985 - shaft reached 1,600 m (1mile); December 1, 1985 - shaft scheduled for completion.

Jan 1, 1985

By Walter E. Dean

Deep-sea manganese nodules are considered to be potential ores of manganese, nickel, cobalt, and copper. Considerable time, effort, and funds have been applied to the study of Se distribution of nodules on the sea floor and the location of economically important areas, but little effort has been spent in understanding the origin of nodules. Most theories that have been proposed to explain the origin of manganese nodules have been based on inorganic physicochemical processes. It is not by coincidence, however, that the area considered to be most economically feasible is in the zone of high biological productivity in the eastern equatorial Pacific Ocean and underlain by siliceous biogenic sediment. The genesis of manganese nodules and the concentration of many trace elements are intimately associated with biological processes.

Jan 1, 1982

By N. C. Miller

In situ stope leaching is an innovative mining system that reduces the surface impacts of conventional underground mining. As in any leaching operation, stope leaching requires hydrologic site characterization and control of leaching solutions. Several tracer tests, using sodium chloride (NaCl) in various concentrations, were conducted by the U. S. Bureau of Mines (USBM) in an underground experimental stope to simulate a leaching environment. The tests were conducted in fractured crystalline rock that was partially saturated. As tests progressed, the surrounding fractures became more saturated. A predominant flow direction away from the stope was found during testing. Testing procedures, preliminary results, and remediation ideas based on field observations are presented in this paper.

Jan 1, 1995

An AFM has been applied to study in situ the adsorption of oleic acid on fluorite in aqueous solutions. The AFM images show that fluorite surface becomes smoother with the adsorption of 7 × 10-5 M oleic acid solution at pH 9. Force measurement shows that fluorite is strongly positively charged in water at pH 6, suggesting a neglectable surface carbonation of fluorite. The adsorption of oleic acid at pH 9 renders fluorite more and more negatively charged. The surface morphology change of fluorite as shown by the AFM images suggests that oleic acid chemisorbs on fluorite at pH 9. Increasing temperature can help increase the adsorption of oleic acid on fluorite. It may also improve the lateral polymerization of unsaturated oleic acid on fluorite as suggested by the smoother mineral surface in oleic acid solutions at an elevated temperature.

Feb 23, 2014

By T. Negeri

Flotation is the most efficient and preferred solid-solid separation method in mineral processing. However, enhanced production of high-grade concentrates at sufficiently high recovery still poses technological challenges to-day. A novel flotation method to produce final-grade concentrates from finely-ground ores in a single step is being developed and tested by CANMET. For the complex sulphide ores tested, grades are superior while recovery is considerably improved. Furthermore, concentrate production rates, at similar recoveries, are faster than for the conventional rougher-scavenger-cleaner configuration. This paper discusses the physical operating parameters affecting Cu-Zn flotation metallurgy.

Jan 1, 1995

By W. G. Pariseau

Recent advances in rock mass mechanics and computational technology now allow for stability analysis of excavations in "wet" mines where mining-induced changes in rock mass deformation and fluid flow are coupled and time-dependent. All the usual dry mine parameters including the premining stress state and rock mass properties are taken into account, in addition to water pressure and flow in jointed, possibly porous, rock masses. Finite element analyses of shafts and tunnels as circular, elliptical, rectangular and arched-rectangular hole shapes excavated in an initially stressed and pressurized rock mass indicate that the usual guidelines for optimum shape and orientation of single openings apply. Here, optimum means maximizing the minimum wall safety factor, or equivalently, minimizing the peak stress concentration. The results also show that failure originates in elements at the wall and that the minimum safety factor at the wall in the wet case is always less than the dry case regardless of drainage time allowed.

Jan 1, 1995

By Milton E. Jennings

On May 11, 1977, the SAMARCO 20-inch, 396-kilometer (246-mile) Pipeline was placed in service. This system represents the latest thinking in slurry pipeline design and is one of the main contributors to the viability of the SAMARCO Project. This paper presents the background, description and latest information on the system's operation. Special attention is given to the handling limitations of the slurry and to the unique hardware requirements.

Jan 1, 1978

By Joel K. Waterland

The Homestake Mining Company at Lead, South Dakota is one of the few remaining gold properties still in operation, in this country. The mine is relatively deep, with a bottom production level of 4250 feet, and a bottom development of 6200 feet. 4, 800 tons of ore are produced daily from three rather large, but inconsistent ore bodies; the dip of which is easterly at about 72 degrees, and the plunge to the south is about 30 degrees. The ore is, for the most part, a chlorite schist with large veins of secondary quartz and cummingtonite schist The ore is not continuous in any direction, but is localized at the ridges of the very intricate system of anti clinal folds. This property has been in active production since its establishment in 1876, except for a period during World War II. The incentive, or contract pay system, has been a gradual process that has been developed by the trial-and-error method down through the years. In looking over the old records, hand-shoveling was paid on an incentive basis as far back as 1916. Since this time, the contract system has grown until it now includes almost all types of underground work - except haulage and maintenance work

Jan 1, 1960

By Michael Prince

"Managing slurry is a constant challenge for mining operations. Since slurries are not uniform, the challenge is complicated by varying rheology, velocity, temperature and pressure, all of which affect how the slurry moves through the pipeline and how destructive it is to the system it is flowing through. Picking the right equipment from the outset — the right size pipe, the right size pump and the proper valves for the application — allows mines to operate more efficiently. Better valve designs deliver multiple benefits — they can lower the cost of operations, reduce the amount of maintenance and associated downtime required and are better for the environment. Historically, modulating systems have used ball, plug and pinch valves to manage the flow of slurry, but each of these alternatives has limitations. Ball valves require ceramic lining in order to delay wear, and once a ball valve does wear, the entire valve must be replaced. Similarly, plugs in plug valves should be coated to perform in control systems, and whole valves must be swapped out upon wear. Pinch valves are expensive and only require replacement of the sleeve. However, sleeve replacement involves an extremely arduous process in which the valve is completely taken apart.Traditionally, designers have avoided using knife gate valves (KGV) in modulating systems because KGVs have not been able to hold up under the harsh operating conditions. The throttling that is required in some of these processes rapidly fatigues conventional KGVs, rendering them useless and in need of replacement. Fortunately, innovative design changes have the potential to change established views.Breaking new groundRecognizing the need for improved components, Victaulic took on the challenge of designing a valve that could withstand abrasive slurry systems for a much longer period of time. The first step in this process was to talk to miners, maintenance personnel, process engineers and operators in the field about the limitations of current offerings and then determine design specifications for an ideal replacement.Many of the responses to these inquiries were very similar and reflected challenges with current operations; the primary objective was to find a valve that would be lighter and cheaper than the valves currently in use. In theory, a lighter valve would be easier to replace, and since replacements need to be performed frequently, a cheaper valve seemed like the right idea."

Feb 1, 2019