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By Renaud Adams, Vladimir Salatic

"The El Mochito Mine located in Honduras, Central America, has produced lead, zinc and silver in concentrates since the 1950s. The paper discusses changes made in both lead and zinc circuits to reduce costs while maintaining production during the recent period of low metal prices.The intent of this paper is to show that sometimes thinking “outside the box” is beneficial for progress. Also, in striving for better metallurgical results we should not forget the basics of flotation. Sometimes the latest trends may not be applicable to the ore we are treating and may not meet the operating cost constraints.A major beneficial impact on El Mochito metallurgy was achieved when the overall grind was coarsened, eliminating the significant percentage of slimes. Switching from two parallel twostage grinding circuits to one circuit with three-stage series grinding has resulted in a significant reduction in power and steel consumption. With minor layout/reagent modifications, both lead and zinc circuits gained not only in retention time but as well in producing better grade products.Of course, the challenges are still there; they exist and only with the full support of operational data collection and metallurgical laboratory investigation success can come. That is why, we as professionals, can never stop in trying to make the most of what we have.INTRODUCTIONThe El Mochito Mine is owned by American Pacific Honduras S.A. de C.V. (AMPAC) a subsidiary of Breakwater Resources Ltd. Of Toronto, Canada. The mine is located in western Honduras, approximately 220 kilometers northwest of the capital city, Tegucigalpa and 90 kilometers south of the country’s second largest city, San Pedro Sula. The mine began operating in 1948 and to date has produced more than 19 million tonnes grading 240 g/t Ag, 3.5% Pb and 7.4% Zn. Production in 2005 was 700,190 tonnes grading 87 g/t Ag, 1.85% Pb and 6.61% Zn.Decreasing metal content in the ore followed by the lowest zinc price on the market in the last 15 years (see Figure 1) has reduced the cash flow and endangered to stop the operation."

Jan 1, 2007

By G. M. Ritcey, K. E. Haque, S. H. Lucas

"Laboratory-scale hydrochloric acid leach tests were conducted on an Elliot Lake uranium ore. Hydrochloric acid leaches without an oxidant or with molecular oxygen (i.e., air or pure oxygen gas) as an oxidant favoured uranium and radium extractions. Conditions for maximum uranium, thorium and radium extractions with hydrochloric acid in the presence of sodium chlorate were determined. The highest extractions of uranium (96%), thorium (81%) and radium (91%) were obtained using a mixture of 44.0 kg HCI and 2.5 kg NaCIOJ per tonne of ore at 75°C with 18 hours ' retention.Flowsheets for two-stage leaching are proposed. The residues obtained from the second-stage leach contain uranium (0.002%), thorium (0.006%) and radium (15 to 20 pCi/g). Tailings of this nature are considered suitable for mine backfill, and, if pyrite is removed, suitable for surface disposal.IntroductionThe nuclear energy industry forecasts an increasing demand for uranium. This increasing demand requires not only largescale mining and processing of conventional and complex uranium ores, but also leads to the consideration of new and nonconventional sources of uranium, e.g., black shales, phosphate rocks, granite, sea water and fossil-fuel ash."

Jan 1, 1980

"Automatic flotation control in varying degrees of sophistication is practiced at a number of concentrators throughout the world. Practically, the variable conditions in the flotation circuit and the difficulty in maintaining accurate instrumentation information all add to the complexity of automatic control.Further development at the Kidd Creek concentrator has resulted in a 3075 mtpd copper flotation circuit capable of running twenty-four hours a day under the direct control of an optimizing supervisory computer.Three main streams, continuously monitored by an XRF analyser, supply circuit information to the optimizing computer. Circuit efficiency calculations (from smelter contract terms) and target setpoints allow the computer to determine the economic optimum direction to move the circuit based on the grade/recovery relationship. Circuit manipulation is accomplished by precise pumping of collector and frother reaqents together with accurate control of flotation bank pulp levels. The former uses computer controlled diaphragm pumps and the latter is achieved by control loops utilizing a float type level transmitter, variable position dart plugs and a process computer working to setpoints determined by the optimizing computer.The development of a fully automated flotation circuit is by no means complete, but increased knowledge and major hardware improvements have allowed its progression to a higher level of sophistication."

Jan 1, 1983

By P. Hajiani

"It has been shown that the economics of the hydrometallurgy of a rare earth process depend upon the amount of acid needed to extract the metal per unit mass of the ore. In this paper, we show how to save a significant amount of acid treating a low grade ferrocarbonatite and extracting +99% constituent REEs. In order to save the acid, the alkali earth metals, originated from gangue minerals were removed prior to REE leaching. Moreover, the operating conditions of leaching reactor are optimized to leave iron behind and to leach REE selectively and effectively. Consequently, the hydrochloric acid consumption per ton of ore drops significantly compared to other processes.INTRODUCTIONThe viability of an extractive metallurgy process of rare earth elements (REEs) to produce an intermediate mixed concentrate depends upon a low-cost simple flowsheet (Bogner, 2014). It has been known that regardless of the mineral, the acid consumption is intimately associated with the operational cost (OPEX) of any REE project, particularly when the initial REE grade is low. Besides acid, the other reagents consumed in the process impact the OPEX. This makes minimizing acid consumption in an REE method a critical priority. The acid consumption in different projects has been reported to be equivalent of 100 kg to 1000 kg HCl (100%) per tonne of ore (Average: 500 kg/t HCl 100%) (Bogner, 2014). This value is strongly correlated with the grade of REE in the ore and therefore in the pre-concentrate.Previously, it has been stated that the Montviel flotation concentrate requires about 300 kg HCl (100%)/t ore to attain +99% REE recoveries using direct leaching (Yu, Verbaan, Pearse and Britt 2013). Hence, for an ore feed rate of 2500 t/d ROM, a large acid production plant (744 t/d HCl 100%) would be required for acid leaching of such concentrate. The equivalent Chlor-alkali plant is estimated to cost M$ 150 (CAPEX) and may require 100 MW power. Therefore, there is a necessity to improve the extractive hydrometallurgy and to reduce the acid consumption significantly."

Jan 1, 2016

By K. V. Konigsmann

"Mattagami Lake Mines Ltd operate a zinc-copper mine with an annual output of about 200,000 tons of zinc and 30,000 tons of copper concentrates. Zinc concentrates arc shipped to a company owned smelter in Valleyfield (a distance of 690 miles), for Export to Quebec City and also to Josephtown, Pennsylvania (1,060 miles, Figure 1.)The protection of concentrates in open gondola rail shipments has been a matter of concern since operations commenced. Considerable financial losses did occur at times because of wind erosion. Operating difficulties were also encountered by customers because of excessive moisture accummulation in the concentrates. Test shipments made in 1969 established the magnitude of losses in open cars. The results and subsequent investigations at our mines and elsewhere confirmed that load protection is an economic necessity.Many types of covers were experimented with, more or less successfully. In 1972 we agreed on a joint program with Dupont of Canada Limited. A new type of cover was developed by early 1973. It has been in use since, undergoing several improvements as experience accumulates."

Jan 1, 1975

By Ben Murphy, Tomi Maksimainen, Antti Rinne, Ishmael Muzinda, Toni Mattsson

"A 500m3 flotation cell was commissioned during 2014 at the Kevitsa Mine in northern Finland working as the first rougher in the copper flotation circuit. Treating around 850tph with an average grade of 0.3% Cu, the TankCell® e500 has run with an average copper recovery of 71% and an average enrichment ratio of 50 during early optimization test work. In addition the cell has proved to be very energy efficient, being operating with a specific energy of around 0.4kW/m3 under production conditions.A review of the installation and operation of the TankCell® e500 at Kevitsa will be presented. Metallurgical performance of the cell during the operating period will be discussed on the basis of on-line analysis and the metallurgical samples collected during early optimization campaigns.IntroductionThe past three decades have seen a rapid increase in the size of mechanical flotation cells. This has been largely driven by demand from the copper industry to process larger volumes of material due to combination of declining head grades and the desire to benefit from the economies of scale (Lynch, Harbort and Nelson 2010). During this period Outotec has been at the forefront of development of large flotation technology with an enviable scale-up track record; 100m3 (1995), 160m3 (1997), 200m3 (2002), 300m3 (2007), 500m3 (2012) to the largest commercially available cell at 630m3 in 2013. This scale-up methodology has been developed, tested and validated (Elgueta et al., 2009; Coleman & Dixon, 2010) by the large installation base of cells in the 300m3 size class operating (over 400). The learning from large cell operation are then reviewed and incorporated back into new designs and scale-up practices (Murphy, Miettinen, and Yanez, 2014)."

Jan 1, 2016

By G. I. Mathieu, R. W. Bruce

Flotation schemes for the separation of molybdenite from other minerals are based on utilizing the Inherent hydrophobicity of molygdenite (Fuerstenau1). This natural floatability was attributed to the crystalline structure of molybdenite by Gaudin et al.2 who suggested that any surface formed by the rupture of van der Wall's bonds should be hydrophobic and, therefore, naturally floatable. Other authors argued that "natural floatability" is dependent of adsorption of hydrocarbon ions and/or molecu1es on the surface of the minerals. This theory, albeit-c questioned 1,2 has long been recognized in actual practice of molybdenite flotation as the latter in carried out with hydrocarbons, such as fuel oil, kerosene and stove oil. Although conferring ready floatability to molybdenite, these presents generally have little effect on other sulphides and silicates. However, talc is a noteworthy exception because of its innate floatability4. Not withstanding the reason for this behavior which has also been attributed to crystal structure2 its end results as in the case of molybdenite is an immediate flotation response to frothing oils arid kerosene, as was observed by Baranovski5, Sutherland and Wark6.

Jan 1, 1974

Westwood Quebec Long Life, Low Cost CONSTRUCTION UPDATE ? Shaft advanced to 840m (after 1 year), 3 km of down ramp and 2 km of exploration drift ? Head frame and hoist room structure completed ? 2009, 84,700m drilling program completed (increase 9%) 2010 OUTLOOK ? $102 million expenditures planned ? Shaft to advance to 1,300m and the first loading excavate 71,500m drilling program planned

May 1, 2010

By H. R. Bullis

"Abstract- This paper discusses what appears to be a flaw in the current belief, widely held within the gold mining industry and investment community, that the creation of ever-larger gold mining companies is desirable for mining companies and investors alike. The basis of the flaw is in the underlying assumption that very large gold producers can discover or acquire new gold deposits of the size necessary to replace extracted reserves on a year-on-year basis over the intermediate to long term (i.e., five to ten years and longer). As it will be noted in this paper, deposits of the size required to replace annual production of very large gold producers (VLGPs) are relatively few in number. Although discoveries of new gold deposits continue to be made, they are predominantly in the 0.5 to 2.0 million ounce range. Data represented here suggest that, although gold production from individual VLGPs continues to increase, the change is due to merging or acquiring other companies or projects with similar reserve life profiles. Therefore, the reserve life profiles of VLGPs have remained flat or have decreased.The paper will note recent changes in the industry and the remarkable increase in gold production reported by individual mining companies over the past several years. It will provide an overview of the geographical and geological distribution of known gold deposits and the range of deposit sizes. Finally, the paper will discuss the challenges that face large gold producers in replenishing their reserves that are being depleted at an ever-increasing rate. Unless otherwise stated, all production and reserve data have been taken from company documents in the public domain or publicly available information."

Jan 1, 2001

By J. Lee, R. G. Rosehart

"The basic problem of arsenic contamination from mine water is discussed. Effective methods of removal involving chemical precipitation, activated carbon adsorption, ion exchange and reverse osmosis are discussed. As well as an economic analysis, detail is also paid to other ions in solution which are removed simultaneously.INTRODUCTIONThe object of this paper is to review arsenic and its removal from gold extraction plant effluents. The first method considered is chemical precipitation. A second method is sorption processes, such as ion exchange resins and activated carbon adsorption. A third method is the possible application of reverse osmosis for the treatment of mine waste waters containing arsenic.Arsenic pollution in the gold mining communities of Canada and Ghana have been extensively studies (l, 2). Lynch(3) Shabunin, Gutman and Kogan( 4), Hollo, Troth, Vennes and Illinigi (5) and Hau(6) have reported their removal methods of arsenic by chemical precipitation, ion exchange and biological methods. Their conclusions were that the precipitation with Fe2+ and Fe3+ ions was most effective.Scope of the ProblemInitial aspects of this research program involved a detailed analysis of the operation of a gold extraction plant located in Canada. The name and location of the mine is proprietary information. Figure 1 gives a basic process schematic with the main tailing streams identified.Table 1 gives information as to the various mine streams and their physical and chemical composition. In this study, primary attention is paid to the removal of dissolved arsenic from waste waters. Solids from this type of operation can always be back-filled. As could be anticipated, the dissolved As content of the primary flotation tailing stream is very low. The dissolved As level in the secondary processing stream and the roasting quench water is quite high. The reason being that at this stage of the process, the arsenic has been freed from the arsenopyrite forms."

Jan 1, 1972

By Don Peel

Mineral aggregate, the most widely used mineral resource globally, has attracted little academic or public interest. Of all mining developments, the aggregate resource is seldom factored into resource management decisions. This lack of recognition to the strategic aggregate resource has significant economic, environmental, and social implications. Impact to the economics, the environment, and the community coincide with the key elements of the United Nations concept of ?sustainable development?. To build on the concept, the mining sector contracted the International Institute for Environment and Development to assess how mining activity contributes to sustainability. As a result, a useful assessment process was developed by the Mining, Minerals and Sustainable Development (MMSD) Task Group entitled ?Seven Questions to Sustainability: How to Assess the Contribution of Mining and Mineral Activities.? In Alberta, recent opposition to mineral resource development by community residents offers an opportunity to employ the MMSD?s assessment procedure. The public opposition to the mining activity was addressed through a stakeholder planning process initiated by the municipal council of Sturgeon County, Alberta. As a result the Calahoo/Villeneuve Area Structure Plan (CASP) was formulated and adopted. To measure the effectiveness of the CASP planning process in terms of sustainability, this paper applies the MMSD?s Seven Question assessment to the Calahoo/Villeneuve Area Structure Plan.

May 1, 2004

By R. G. Hamm

A brief outline covering the shovels and draglines in use in the open-pit mines of Western Canada, including British Columbia, Alberta and the Yukon, is given, followed by the procedure for the selection of the correct wire rope for each application. Such fa ctors as wire rope constructions, lay directions and types of cores are mentioned. Once the correct wire rope has been selected, there are variousfactors that can affect the service fife ofthe rope and the reduction ofdown-time during rope change-overs. Factors such as hairpin winding, types of Becket loops, and shovel drum and bail adaptors are described.

Jan 1, 1979

By A. P. E. Dirige

Extensive field study has been performed to assess the capabilities of thin sprayon liners (TSLs) and conventional spray-on support systems for preventing rock and support material damage that often results due to mining-induced rockbursts. Support performance was studied using field scale explosive detonation trials to simulate dynamic failure effects that are known to develop during typical rockburst events. Multiple seismic and high speed photographic monitoring techniques were used to provide detailed information concerning rock motion, surface fracturing, ejected fragment motion, and support liner survivability characteristics. Results of this study have demonstrated that TSL?s and variant layer combinations may be as effective as or better than conventional support materials for mitigating rockburst or like damage in highly stressed mine environments. In this study, field results have been verified using numerical modeling procedures to better understand the support behaviour of TSLs when subjected to highly stressed mine environments and mining-induced rockbursts, a factor that is essential when designing any rock support system. A series of numerical modeling assessments were conducted using FLAC3D modeling of a half circular TSL-lined tunnel, influenced by anisotropic stresses and simulated mining-induced rockbursts, to investigate the support potential of TSLs under prototype conditions. The effect of a mine rockburst was simulated using the three-dimensional dynamic analysis option of FLAC3D. Results of TSL-lined tunnels in highly stressed and mine rockburst conditions indicated exceptional effectiveness of TSLs for suppressing rock deterioration resulting from stresses and rockbursts. The rock support potential of TSLs on tunnel surfaces was compared with that of thin shotcrete linings, and results indicated that thin spray-on lining products currently available may be equally as effective in support as shotcrete materials.

May 1, 2009

By Ish Grewal

"Batch centrifugal gravity concentrators, which produce very high ratios of concentration, have become common place for recovering free gold within milling circuits. Their effectiveness has proven to be especially useful in recovering finer gold particles. The knowledge related to testing, modeling and scale-up of these systems has been well established over the recent decade mostly due to the efforts of the late Dr. Andre Laplante.Gravity separation technologies such as jigs, spirals and shaking tables can provide the increased mass yields required for minerals that occur at higher concentrations than gold, however their effectiveness tends to diminish at finer particle sizes. The development of high mass yield centrifugal concentrators has provided the opportunity to be able to recover a wide range of minerals at finer particle size classes when compared to the more traditional gravity concentration devices. These concentrators also provide additional benefits such as smaller unit footprint and automated operation.This paper discusses the background, current practices and future applications for high mass yield centrifugal concentrators in base metals, sulfides and industrial minerals. Bench scale, pilot scale and plant practice examples are also presented.INTRODUCTIONGravity concentration, one of the oldest mineral beneficiation processes, has evolved over the decades and continues to play an important role in modern mineral processing operations. The development and growth of froth flotation in the early part of the twentieth century led to a decline in the importance of gravity concentration. However, gravity concentration continued to be used for ores that are difficult to beneficiate using froth flotation; such as tin, tungsten, niobium/tantalum, iron, industrial minerals and oxidized coal. In recent decades, many operations have begun to employ a combination of both gravity concentration and froth flotation Since flotation relies on surface chemistry and gravity concentration relies on specific gravity differences, putting these techniques in series provides today’s metallurgist with unprecedented opportunities to maximize grade and recovery of many of these minerals."

Jan 1, 2009

By Hongfei Li, JianLong Wang

This paper introduces the technical characteristics, transformation scheme and production practice of the project in which Copper P-S Converter was replaced by Bottom Blowing Continuous converting for the first time, and introduces the technical achievements obtained by the project. The application practice proves that bottom blowing copper continuous converting technology can completely solve the problems of fugitive off-gas pollution in the ladle transportation, large amount of flue gas and fluctuation of flue gas caused by P-S converting. Comprehensive energy consumption and operating costs of producing blister copper, safety and environmental risks are greatly reduced, and economic, environmental and social benefits are significant. It is one of the preferred technologies for upgrade and to replace P-S converting technology. INTRODUCTION At present, more than 80% of copper production capacity is still using the P-S converting technology with a history of more than 100 years. During P-S converting process, melted copper matte needs to be transported by ladle, and it needs to rotate the furnace body several times, causing a lot of serious environmental and safety risks such as SO2 flue gas emission and low-altitude SO2 pollution. In addition, the high operating cost caused by low rate of air blowing, the short life of the refractory materials and high energy consumption during P-S converting become big problems. Other problems such as large amount of flue gas, low SO2 concentration, large fluctuation of flue gas volume and SO2 concentration, and discontinuous flue gas cause high cost of subsequent flue gas treatment are also needing to be solved. In order to solve these problems during the P-S converting, metallurgical engineer in various countries has spared a lot of efforts to develop continuous blowing process for decades. The continuous converting that have been industrially applied are flash continuous converting, Mitsubishi continuous converting, Noranda continuous converting, Bottom Blowing Continuous converting and multi-lance top blowing continuous converting

Jan 1, 2019

By Gerardo Vargas

Without an adequate data sample there can be no Reliability Analysis (RA). Without analysis there can be no systematic, verifiable improvement in reliability or in operational economy. A sample is a collection of life cycles. A beginning and ending event define a failure mode life cycle. Well known obstacles impede the reliability engineer in his role to analyze maintenance data. The main problem lies in the difficulty to obtain analyzable data samples. Although modern Computerized Maintenance Management Systems (CMMSs) should provide the needed information they rarely deliver samples of adequate quality. Cerrejon, an integrated mining operation in Colombia, South America solved the information management problem by applying a "Living Reliability Centered Maintenance (LRCM)" process to its fleets of Trucks, Graders, Dozers, Loaders, and Shovels. This paper describes a method wherein completed work orders capture reliability analysis enabling information. The 'right' maintenance observations reference significant failure modes. Work orders link to records in a continuously growing Reliability Centered Maintenance (RCM) structured knowledge base. Grouped and filtered knowledge-to-work order links, being instances of failure modes, provide the samples required by reliability analysis software tools and methods. LRCM software and related procedures facilitate the growth of knowledge, manage the work order-to-RCM relationship, and generate samples for subsequent reliability analysis.

May 1, 2011

By R. E. Jones

"Crusher sizing and application in mining operations is the selection of a crusher or a series of crushers to reduce run of mine ore to a size suitable for either autogenous or conventional grinding. It is important that the crusher or crushers be sized to operate within the range to which they have been designed. By the range to which they have been designed, we mean that the reduction ration must be held between specified limits. Reduction ration is the relationship between feed size to product size.In todays plants reduction is usually accomplished in three stages by primary, secondary and tertiary crushers.Primary crushers are either jaw or gyratory and are designed to handle run of mine feed at an open side setting of 5-1/2” to 7”. Reduction ration should not exceed 7 to 1.Jaw CrushersThe model size of a jaw crusher signifies the feed opening. A model 60 x48 jaw has a feed opening 60” wide by 48” long. The 48” is the dimension between the top of the swing jaw and the top of the stationary jaw. The 60” is the width between the two side frames. The crusher open side setting is the discharge opening at the bottom of the crushing chamber with the jaws in an open position. When corrugated jaw plates are used, the open side setting is measured from tip of one jaw plate to the valley of the opposite jaw plate. If a jaw crusher has a 6” open side setting with a 1” jaw motion, you could also say it has a 5” closed side setting."

Jan 1, 1974

By Paul H. Atkinson

"Sampling in the mineral processing industry has always been a subject which has received considerable discussion. There have been many volumes written on theory, techniques and practical application of sampling. However, until recently, the companies were not completely aware of the potential savings available with properly designed sampling systems both to assist in pieces: control and final product values.Definition In our discussion today, we will define sampling as the operation of removing a part, convenient in quantity for analysis, from a whole, which is of much greater hulk, and removing it in such a way that the proportion and distribution of the quality to be tested, i.e. size of particle, moisture content, specific gravity content, mineral and chemical composition, are the same in both the sample and the whole.Application It has been well established that undetected errors can result in large economic losses, thus most firms depend upon some type of sampling and sample analysis as a control in their continuous processing of materials. Despite the large investment in facilities and personnel the ""analysis"" of a sample can be no more accurate or reliable than the sample on which it is being based. The importance of sampling is recognized by a great many but it is appreciated by only a few, largely because it is sometimes one of the items which is taken for granted."

Jan 1, 1976

By Seppo Haapamaki

"Since it started production in 1970, Anderson Lake Mine has experienced several radical changes in mining methods as a result of the physical characteristics of the ore zone along with the need for improved productivity. This paper examines the transition from longhole open stoping to mechanized cut-and fill then to vertical crater retreat at Anderson Lake Mine. The reasons for the need to change mining methods is analyzed taking into consideration the impact of such criteria as improved productivity and ground conditions and how these criteria affected the over-all economic viability of the mine. IntroductionAnderson Lake Mine is one of five producing copper mines in the Snow Lake area and is situated approximately 3 km southeast of the townsite of Snow Lake (Fig. I).The mine production rate at Anderson Lake is 700 tpd and all ore is trucked from the mine to the mill located approximately 2.5 km from the mine site . The mine is serviced by a three-compartment shaft sunk to a depth of 3300 ft with level spacings at 175 ft and 150 ft intervals from the 600 ft level to the 3000 ft level. .The shaft was collared in 1965 and, a year later, the hoist and compressors were installed. By 1969, a crusher system had been built and actual production began in 1970. During 1978, production was suspended for the year to allow for 430 ft of shaft deepening."

Jan 1, 1988

By X. Louw

"This paper examines the change in surface properties of the Polyurea used on spirals at different operational times: unused spiral, 28 months in operation and 58 months in operation. This is being done to investigate the alleged statement that spirals have a decrease in spiral separation efficiency after a certain period of operational time. Observing the surface characteristics under 1.5x and 36x magnification it can be concluded that the surface roughness decreases with increased operational time. This may be explained by a decrease in air inclusions and streak lines as the spiral surface wears. Decreased surface roughness should in theory have a beneficial effect on spiral separation performance. This leads to the conclusion that if a decrease in spiral separation performance is observed after a certain period of operation, it can’t be attributed to the change in surface properties. Uneven wear of the spiral surface, material creep or application of force to the spiral will change the profile of the spiral and could cause an adverse effect on spiral separation performance. The metallurgical test work showed that the best results are obtained at 30 - 35% solids by mass in the spiral feed. The iron recovery was significantly higher in the new spiral in comparison to the 2012 and 2010 spirals. The mass yield to concentrate in the new spiral was approximately three times more than for the older spirals. In conclusion, the yield combined with the grade, which equals the recovery was the best for the new spiral.IntroductionAn investigation was launched to research the common belief that spirals experience a decrease in efficiency, and therefore recovery, after a certain period of operation (a period of seven years has been used in the industry). The properties of the spiral that may have an effect on the efficiency are:"

Jan 1, 2017