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By Y. Lahaye, O. Martinsson, B. I. Palsson

"Four main and three minor rare-earth-element (REE)-bearing minerals were identified and quantified in the Kiirunavaara apatite iron ore tailings using optical microscopy, an electron probe microanalyzer (EPMA) and a mineral liberation analyzer, and their chemical compositions were analyzed by the EPMA and laser ablation inductively coupled plasmamass spectrometry. REEs are shown to be contained in the minerals apatite, monazite, allanite, titanite, zircon, thorite and synchysite. In zircon, thorite and synchysite, they occurred in only trace amounts and contributed limited amounts to the total REE budget, and these are consequently of minor importance. Monazite occurred as inclusions in apatite and as free particles, 90 percent liberated. Allanite occurred to some degree in mixed grains with magnetite but also as free particles. Monazite mainly reported to the apatite concentrate, while allanite and titanite largely went to the tailings, the latter preferably to those fractions smaller than 38 µm. The amount of titanite in the finest tailings fraction was 2.3 weight percent, containing close to 1 percent REEs, with heavy rare earth elements (HREEs) making up 28 percent of the total REEs. However, a texturally distinct group of titanite grains showed an HREE/REE ratio of up to 67 percent. Furthermore, titanum dioxide analyses indicate that titanite is preferentially released into the tailings from the secondary magnetic separation step in the concentrator. Our data therefore suggest that titanite, occasionally enriched in HREEs, can be extracted from the processing stream and might thus be considered a new source for REEs at Kiirunavaara and similar deposits.IntroductionDeposits of apatite iron ores are the major source for iron in the European Union and are exploited by Luossavaara-Kiirunavaara AB (LKAB) in several mines in the Kiruna-Malmberget area in northern Sweden. Kiirunavaara is the largest of the apatite iron orebodies in Sweden, comprising more than 2 Gt of iron ore with 60 to 68 percent iron (Fe). At the end of 2015, the total accumulated production of crude ore from openpit and underground mining since the start of mining was 1,566 Mt, with ore reserves of 644 Mt grading 46.3 percent Fe (LKAB, 2016). The deposit defines a steeply dipping tabular body, approximately 6 km long and up to 180 m thick, extending at least 1,500 m below the surface. It is situated in the middle part of the Palaeoproterozoic Kiirunavaara Group of volcanic rocks and follows the contact between a thick pile of trachyandesitic lava and overlying pyroclastics (Martinsson, 2004)."

Jan 1, 2017

By R. Costa, A. Kozak, B. Fukuhara

Introduction The Jacobina mine is located in the state of Bahia in northeastern Brazil approximately 330 km northwest of the state capital city of Salvador and 9 km from the town of Jacobina. The Jacobina Mine (Figure 1) is owned by Jacobina Mineração e Comércio, a wholly owned subsidiary of Yamana Gold Inc. The gold mineralization of the Jacobina mine is hosted within quartz pebble conglomerate reefs of the Serra do Córrego Formation which together with the Rio do Ouro Formation forms the Serra do Jacobina mountain chain. The Serra do Jacobina mountains have been mined for gold since the late 17th century. Since the 1970’s, the Jacobina mined has been operated by Anglo American, Williams Resources Inc., Desert Sun Mining Inc. and now Yamana Gold Inc.

Jan 1, 2007

By R. Scheffel, R. Washnock, G. Zarate

"The second survey of global copper leaching operations has been completed, the data published in 1998 has been updated. A survey questionnaire was sent to 40 identified copper leach operations around the world including those leach properties still in operation that responded to the 1998 survey. This paper lists the results of 26 operations that responded to the survey. Selected information from the survey is represented in graphical form. INTRODUCTION Copper leaching was first performed by truck dumping huge monolithic stockpiles of ore and flooding the surface with leach solution or lixiviant with little understanding to the parameters that support optimum recovery of the metal. Heap leaching has evolved into a controlled process whereby a complex low grade ore is stacked in shallow lifts (under 10 m in most cases). The ore is usually crushed and agglomerated, conveyor stacked on an engineered containment system (leach pad) and irrigated in a controlled manner. The lixiviant chemistry is designed to extract the optimum amount of metal from the ore. At first glance, heap leaching appears to be simple in nature. In fact, it is a very complex process. A plethora of technologies such as material handling, environmental control, analytical techniques, solution chemistry, hydrology and many others contribute to the design and operation of a heap leach. Heap leaching requires an inter-disciplinary approach between geologists, mining engineers and metallurgists. Many operational factors affect heap leaching. The process must address climatic factors, flexible operation, quality control before and after leaching, physical characteristics of the ore, geology, mineralogy, reagent consumption, solution application rate and methods (wobblers or emitters), particle size, and favorable conditions for bacterial growth and survival. Predicting, troubleshooting, and optimization of leach recovery is a significant challenge when considering the number of variables involved and their interactions The survey updates many of these key operating parameters to the current state of the art. It should provide the reader an excellent benchmarking summary of these factors on a site-by-site basis."

Jan 1, 2016

By C. L. Schneider

The calculation of the liberation spectrum in the product from continuous milling operations is a problem of considerable importance and is difficult to solve because the population balance equation must be integrated in two-dimensional space. Complete solutions require the estimation of a large number of parameters unless a structural model of the ore texture can be used. A very simplified textural model based on the use of the modified Andrews-Mika diagram and the PARGEN simulation of multi-phase particles has been formulated and is used to simulate the behavior of a continuously operating laboratory ball mill. A sensitivity analysis for this mill has been done to show the likely effect of model approximations on the predicted liberation spectrum which will be measured.

Jan 1, 1992

By K. Osseo-Asare

LIX63 in combination with DNNS selectively extracts nickel and/or cobalt from acidic leach liquors with pH values as low as 1.0. Nickel/cobalt and cobalt/nickel selectivities are achieved by varying the relative concentrations of LIX63 and DNNS. Increasing LIX63 concentration retards both extraction and strip-ping rates while DNNS has the opposite effect. Slope analyses of equilibrium data indicate that both LIX63 and DNNS are present in the extracted complexes. Examples of possible flowsheets are presented.

Jan 1, 1978

By R. L. Bleifuss

The object of this study was to evaluate the oxidized taconites of the western Mesabi iron range and to establish a correlation between the various basic taconite types and their concentratability. This study was made at the Mines Experiment Station of the University of Minnesota under the sponsorship of the Great Northern Railway Company. The initial work included a study of the original mineralogical and stratigraphic facies of the fresh taconite and traced the oxidation and leaching patterns developed during the conversion of this material to oxidized taconite. Subsequent studies have extended this work to include the semi- taconites and also the wash and heavy media ore types widely developed in the area. Each basic, original mineralogical facies or taconite type produces during alteration an equivalent secondary oxidized facies which has a unique mineralogical composition and texture. These control the concentratability of the oxidized taconites. Actually the influence of the original mineralogical composition of the iron formation extends even further and determines whether the most likely product of oxidation and leaching will be a wash ore or a heavy media ore and even to the direct shipping ores where the ratio of hematite to goethite is largely determined by the original taconite type.

Jan 1, 1964

By J. R. Heberger, B. H. Ngo

The Safety Pays in Mining v2.0 web application, developed by the National Institute for Occupational Safety and Health (NIOSH) Mining Program, helps mines determine the potential costs associated with mining injuries. This web app categorizes injury cost by part of body injured, by the cause of the injury, or by the nature of the injury. When the user selects one of over 150 common types of mining injuries, the app provides information on the distribution of costs of workers’ compensation claims for that type of injury. Based on other user inputs, Safety Pays in Mining v2.0 will estimate the total costs of the selected injuries and the estimated impact of total injury costs on mining company profits and will provide some examples of services and personal protective equipment on which companies could spend the savings that result from the prevention of injuries. This paper reviews the Safety Pays in Mining version 2.0 web application by discussing the development and updates to the app, how it is used to show the true costs of mining injuries, and how mines can benefit from using this app.

Feb 1, 2024

By S. S. Sephton

Sintering as applied to agglomeration of ferrous fines has shown little growth in the United States in recent years. The general concensus was that most of the existing facilities would remain in production at a reduced tonnage to dispose of and recover solid waste from steel plants. Further growth for ferrous sintering technology was not in sight in the United States and in the early sixties expansion came to a virtual standstill, however, we witnessed a rapid emergence of a tremendous pelletizing industry. On other continents, due to different raw materials supply and capital structure, sintering technology continued to grow, with a more cautious development of pelletizing installations. In the U.S. the diminishing supply of good quality, high grade coarse ore resulted in eventual disappearance of high grade ore fines for sintering. The finely divided concentrates produced from lower grade ores were not amenable to this type of agglomeration, but were more suitable for pelletizing. Furthermore, pellets seemed to possess all the features of an ideal blast furnace feed; namely pellets are free flowing, uniform in size, strong, with extremely uniform chemical composition and good reducibility. Importance and demand for such well stabilized furnace feed further increased as size of newly installed blast furnaces continued to grow. The in- creasing demand for a uniform, stabilized burden coincided with the effort to produce a fully beneficiated burden for smelting, thus reverting the blast furnace to its primary function of smelting.

Jan 1, 1973

By K. Perry, B. Lusk, Q. Huang, R. Honaker

"Rock dust is applied in underground coal mines to prevent the propagation of coal dust explosions. The application is conducted either dry or wet. Both methods have advantages and disadvantages however the formation of caked rock dust particles is cited as a significant concern for wet applications. A study was conducted to evaluate the effect of modifying the rock dust particle surfaces to repel water and each other by the addition of oleic acid (OA) and sodium oleate (NaOL). At a NaOL dosage of 2 lbs/ton, a contact angle of 112o was measured indicating strong surface hydrophobicity and ability to repel water. FTIR studies indicated a chemisorption mechanism which is necessary for long term stability of the surface modification. The dispersion rate of the modified rock dust samples was elevated by the application of both OA and NaOL thereby indicating a significant reduction or elimination of the detrimental effect of caking. Tests in an explosion chamber showed a 79% reduction in the dust explosion potential relative to untreated rock test applied by the wet technique. INTRODUCTION Mine explosions have been a long-time threat to underground coal miners attracting significant attention to the accumulations of combustible dusts and gasses. In the presence of sufficient oxygen and heat, the accumulations of combustible dust and gass can lead to large-scale underground coal mine explosions [Harris et al., 2010]. To alleviate the problem, rock dust such as limestone can serve both as a thermal inhibitor and a heat sink to inert the coal dust explosibility thereby prohibiting the occurrence of dust explosions [Richmond et al., 1975; Amyotte et al., 1995; Dastidar et al., 1997; Man et al., 2009]. The dispersibility of the rock dust is an important characteristic that determines the degree of explosion prevention [Greenwald et al., 1938; Mozumdar et al., 1974; Amyotte et al., 1995]. Unfortunately, the environmental conditions of an underground mine has the potential to cause caking of the rock dust which reduces the dispersive characteristics needed. The dispersive property of rock dust can be enhanced by modification of the particle surface properties by the application of chemical additives [Greenwald et al., 1938; Mozumdar et al., 1974; Spence, 2004; Vogt, 2011; Doidge et al. 2014]. The major mechanism for rock dust caking is the adsorption of water which forms liquid bridges at the particle-particle contact points where saturated solutions of soluble powder ingredients are present. Changes in temperature or relative humidity result in solvent (water) evaporation and the formation of dry bridges [Rumpf, 1958; Specht, 2006]. One method to prevent caking is creating hydrophobic rock dust to repel water from the solid surface thereby preventing the formation of liquid and solid bridges."

Jan 1, 2015

By Charles B. Escapule

The first successful heap leach of silver ores in the State of Arizona has been under continuous operation since early 1976 at the State of Maine Mine, two miles southwest of Tombstone, Arizona. The silver produced from leaching old mine dumps at this property vas the first in the Tombstone mining district to achieve a 999+ purity. While cyanide leaching of silver and gold ores is a time-proven accepted practice, a relatively new method of treatment called "heap" leaching was employed. Heap leaching is a comparatively inexpensive fast way to recover precious metals from low grade material.

Jan 1, 1981

By J. Y. Hwang

Coal cleaning with traditional magnetic separation usually yields unsatisfactory results because minerals associated with coal, such as quartz, clay, pyrite and calcite, are weak-or non-magnetic. Magnetic reagent technology is a new mineral processing technology with the capability to enhance the magnetic property of minerals selectively. This reagent can be adsorbed selectively on minerals through the control of surface chemistry and, thus, result in selective magnetic enhancement. The application of this technology on coal cleaning has been experimentally demonstrated. Good separation results can be obtained at low to medium magnetic field strength. Comparison with other processing technologies is included.

Jan 1, 1992

By S. Tuylu, D. Adiguzel, A. Bascetin, E. G. Savas, O. Ozdemir

"Paste Technology has some advantages comparing to other surface disposal methods. With this method, it is possible to pump high concentrated solid materials to very long distances without segregation problem. However, it is known from Literature that pumpability of materials is completely related to rheological properties of the materials. For this reason, in this study, the rheological parameters (e.g. yield stress and viscosity values) of the paste material prepared from Pb-Zn tailings at various solid ratios (60-85%) were obtained in order to determine the pumpability of the paste material. Additionally, slump tests, which are usually used in mining industry, was carried out. Finally, regression analysis was made between the slump values and rheological parameters of the paste material to find a correlation between two independent parameters. The results showed that there is a great correlation between the slump values and the yield stress of the material. On the other hand, there was no correlation with viscosity: ty=430.28-16.877Sv (R2= 0,94; ty: Yield Stress (Pa), Sv: Slump Value (cm)). Based on the results obtained from this study, it can be clearly concluded that the slump values obtained from a simple test can be used to predict yield stress values of paste materials. INTRODUCTION Recently, surface disposal of mining tailings has become one of the biggest problems in mining industry in terms of environmental issues. Therefore, several researches have been focused on this subject in order to find a best solution or method for this problem. These studies showed that paste technology has some advantages compared to other conventional methods. Tailings thickened to the extent show no segregation of particles during transportation, and relatively small amount of settling occur post-deposition which are often called “paste” (Cincilla et al. 1997). Paste tailing technology has two applications in the industrial area as underground paste fill and surface paste disposal. Paste tailing technology was first started to be used in Grund mine, Germany in the late 1970s (Grice, 1998). Process tailing material, water and - if needed - cement materials, which are the components of the paste, must be blended to obtain a homogenous mixture suitable for pumpability. Consistency behavior of the paste material is determined using a conventional sedimentation cone concrete test (slump). Meanwhile, the slump value of the paste must be between 200 and 250 mm for surface applications (Newman et al., 2001). Additionally, in the case of underground paste system, the optimum slump value must be between 150 and 200 mm. On the other hand, the water content and density of paste at certain slump value change depending on the amount of fine particles in the tailing sample. While moisture of the material will be high, its density will be low because of surface area of particles which increases with decrease in particle size (Meggyes and Debreczeni, 2006). For pumpability of the paste to long distances, slump value of the paste must be good enough to be accepted for use of surface paste disposal. Paste material has a yield stress because they show non-Newtonian behavior and generally is classified as a Bingham plastic fluid. (Cooke, 2008; Theriault, 2003). In addition, the appropriate viscosity value of the paste along with the yield stress must be between 100 and 250 Pa (Bascetin et al., 2013; Bascetin et al., 2014; Bascetin et al., 2015)."

Jan 1, 2016

By Maurice C. Fuerstenau, Bruce R. Palmer, R. E. Johnson, B. R. Palmer, G. W. Warren

The environmentally sound processing of metal-bearing chloride solutions plays a key role in the extraction, plating, recycling, finishing and surface preparation of metals. Additionally, many chloride-bearing waste streams must be processed to obtain environmentally acceptable products. This paper reviews the technologies used for processing aqueous metal-chloride solutions and describes the challenges that have been overcome in application of these technologies.

Jan 1, 1993

By Christopher J. Bise

It is somewhat reassuring to note that the United States has an abundance of coal reserves which can be used to bridge the widening gap between the nation's energy production and consumption. The U. S. Bureau of Mines has recently established a demonstrated coal reserve base of 434 billion tons. (31)(Figure 1) As can be observed, the major percentage of the reserves is in the West. Although surface coal mining in the West has been gaining increased attention in recent years, western reserves that must be deep mined are significantly greater. Though deep mining activity in the west is very limited and accounts for less than 2% of the annual production, the primary reason, quite probably, is the greater thickness of the seams as compared to those found in the East. Forty-five billion tons of the western demonstrated coal reserve base for underground mining are in seams greater than 10 ft. in thickness. (37) Thick coal seams do not lend themselves readily to extraction with the underground mining methods practiced in the United States today. Experiences from the practices in other countries can serve only as guidelines. Consequently, a concerted effort is necessary to overcome the technical difficulties.

Jan 1, 1975

By Nikhil Dhawan, Shrey Agrawal, Ashwini Kumar

The current study investigates sericite clay as an alternative source of potash and alumina. The alumina and potash content in the sericite is approximately 35 and 10%, respectively, compared with 50–55% in bauxite ore. Muscovite and illite phases are stable, and direct leaching is found to be ineffective. Here, planetary ball milling was performed, which provided dissociation of stable phases to relatively reactive phases. The milling led to a five-fold increase in the surface area, resulting in enhanced leaching kinetics even at low acid concentration. The statistical design obtained optimal conditions of 8 h of milling followed by leaching in 1.73 M HCl at 75 °C, yielding 82% alumina extraction in precipitate with 85% purity in γ-alumina form and 63% potassium dissolution in sylvinite phase. The yield of the precipitate was ~65–68%, and the acid leach residue contained primarily silica.

By A. S. C. Owili-eger

An efficient state-of-the-art dynamic model to simulate and predict the rate of water production in longwall coal mining operations was developed as a tool for mine planning purposes. The model can also provide an estimate of any potential resultant hydrogeological impacts of an operation via an array of element-specific single-valued parameters referred to as the field rating indices. The computational procedure was designed to fully exploit the benefits of hydrogeotechnology and utilize a combination of conceptual, physical, empirical, and numerical approaches to define, characterize, analyze, and evaluate flow con- trolling properties of fractured rock systems. The model presented here provides a dynamic prediction method for highly fractured flow; a practical tool in mine and water control planning; and a cost-effective means of assessing, based on the best available technology, the impacts of mining on the water resources within a mining environment to satisfy local and federal regulatory requirements. Unique features include consideration of laminar and non-darcian flow domains in double-porosity-pressure hydrogeological media.

Jan 1, 1987

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

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

Feb 1, 2023

By Omid Asghari, Amin Mousavi, Arshad Ranjbar

Geometallurgical modeling of the production benches is an attempt to estimate the geometallurgical variables of the different ore domains, which leads to a better understanding of the ore and its properties. Having geometallurgical variables results in a more accurate prediction of ore characteristics, an improved drilling and blasting operation, and better efficiency of the processing plant. In this research, the geometallurgical variables required for the geometallurgical modeling of the extraction benches are discussed. Then, the relations between Bond work index (BWi), as one of the most important geometallurgical variables and the primary grindability index, and rock geomechanical properties are investigated based on the previously published studies. The BWi is the main parameter for measuring the resistance to the grinding of ore minerals. It represents the energy required for processing (e.g., crushing, grinding) ore rocks in the beneficiation plant. As a real case study, BWi of a production bench in Sungun Copper Mine, located in East Azarbaijan province, Iran, was estimated. The rock’s geomechanical properties were measured using laboratory tests, including uniaxial compressive strength, Brazilian tensile strength, point load strength, and hardness. Then, BWi was estimated using the proposed formulas in the literature and the results of the rock’s geomechanical properties. The selected bench consists of three ore zones and two dyke domains. The laboratory tests revealed that the first ore zone has a slightly higher compressive strength than the other two zones, while the compressive strength of the waste rocks is larger than the ore domains. Using the proposed equations in the literature, the predicted values of the BWi for the samples in ore zones are in the range of 6.36 – 14.44 (kWh/t), and therefore this wide range must be considered when the short-term production schedule is planned. In addition, variations are observed in the samples within a specific zone and between the ore zones, demonstrating heterogeneity in the selected bench. Finally, wide-open results obtained by different equations show a necessity to perform a geometallurgical program specifically for the Sungun mine.

Oct 28, 2021

By Richard Jr. Ray, Richard Druesedow

The Virginia Operations of Consolidation Coal Co. include a diversity of mine fan configurations which have been tested to quantify pressure losses through the inlet ductwork. Airflow and pressure measurements were taken at four shaft sites to calculate losses for the 90-degree bend at the top of the shaft and the self-closing dampers or closing doors located between the shaft and fan. Velocity contours were then developed from the test data to determine flow patterns in each fan system component. In addition, yaw and pitch angles of airflow through the fan annulus were measured to evaluate turbulence levels created by various inlet ductwork designs. The test results indicate that proper design of the shaft collar and ductwork arrangement for main mine fans can substantially reduce system pressure losses and air turbulence. BY minimizing the pressure drop experienced on the surface, these design improvements enable the bulk of the total pressure developed by the fan(s) to be utilized in the mine.

Jan 1, 1995

By Ivor Hawkes

INTRODUCTION High energy breaking is an alternative to using ex¬plosives in underground secondary breaking operations. It also is a means of upgrading conventional hand-held breakers, manual sledge-hammer breaking, and scaling bar operations. Major areas of application are in sec¬ondary breaking over grizzlies and at drawpoints. Other applications include breaking down ripping lips in longwall seam mining, scaling in stopes and rooms, general demolition work, and roadway maintenance. There is considerable interest in high-energy impact breakers for use in primary ore breaking, but, as of 1977, all such applications have been only experimental (duToit, 1973; Joughin, 1976; Wayment and Grantmyre, 1976). EQUIPMENT Essentially, a high-energy impact hammer is a boom¬mounted pneumatically or hydraulically actuated breaker. The machine basically consists of a piston that oscillates in a housing and impacts the end of a tool or moil thrust against the rock. The force applied to the rock primarily depends upon the impact energy of the piston-the higher the impact or blow energy, the greater the force and, thus, the greater the rock break¬age. Among drill and breaker designers, a common expression for blow energy is "force of blow." Hand-held breakers are limited to blow energies of about 140 J (100 ft-lb), because the operator is unable to handle heavier machines efficiently or to absorb the recoil energy resulting from higher blow energies. How¬ever, these restrictions do not apply to boom mounted breakers; machines with blow energies on the order of 4000 J (3000 ft-lb) and higher are available commer¬cially for underground use. There is considerable evi¬dence to show that increasing the blow energy also in¬creases the efficiency of the breaking operation; i.e., more rock is broken per unit of energy expended (Grantmyre and Hawkes, 1975). Thus, there is a trend to higher blow-energy machines, particularly where high¬strength rocks are to be broken. In relation to rock breaking, the blow rate of boom¬mounted impact breakers is not as important as it is for rock drills. This is because the breaker must be moved over the work surface between blows. The blow rate is governed eventually by the power supply, and typical blow rates range between 200 and 600 blows per minute. As a general rule, light blow-energy machines have higher blow rates than heavier machines. Table 1 lists most of the boom-mounted impact breakers that were available commercially during 1977, and it gives details of the blow energies and machine weights. Restrictions are placed on the blow energy by the machine weight and size, and by the strength of the boom. Typically, boom-mounted impact hammers have a blow-energy to mass ratio of about 1.5, with lower values for lighter machines and higher values for heavier machines. In addition to supporting the hammer weight, the boom also has to absorb the recoil energy of the blow, which can be on the order of 1400 J (1000 ft-lb) for large hammers operating in a horizontal mode. Interesting exceptions to the general run of impactors are the Joy HEFTI hydraulic hammers. In these machines, the piston impacts onto a fluid cushion that is positioned between the piston and the impact tool. This approach allows very high piston velocities, over 30 m/s (100 fps), to be used without the risk of break¬ing the piston or impact tool. Steel on steel impacts must be limited to impact velocities of about 10 m/s (35 fps) due to the high impact stresses generated; thus, increased blow energies can be achieved only by increas¬ing the piston size. The Joy 514 HEFTI®, listed in Table 1, has a blow energy of 27 100 J (20,000 ft-lb), but, as of 1977, the machine has been used underground only on an experimental basis. Using a fluid cushion between the piston and the impact tool allows the use of light pistons, reducing the overall machine weight. The recoil energy, which must be absorbed by the boom for a given blow energy, is directly proportional to the piston to machine mass ratio, and operating with light pistons provides an addi¬tional benefit in reducing the requisite boom size. Both pneumatic and hydraulic hammers are avail¬able commercially. Although hydraulic hammers are a relatively recent development, they already outnumber the pneumatic machines in use. There are many reasons

Jan 1, 1982