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By J. R. Gump

The term "rare earths" is used to designate a group of elements, Numbers 57-71 which are closely related chemically. They were called "earths" because their oxides resemble those of the alkaline earths, and the designation "rare" was originally applied because of their scarcity. Now many of these elements are known to exist in considerable quantities, and the term "Lanthanide" is applied to this group of elements, because of their similarities to lanthanum, the first element in the group. Many authorities insist on including yttrium, element Number 39, because of its extremely close association with the members of the group in nature and because of its similarity in chemical properties. It has been estimated that the combined rare earth group makes up about five-thousandth. of one percent of the Earth's crust, a higher percent than the common elements, lead, tin, zinc, arsenic, mercury, silver and gold. Cerium is by far the most abundant member comprising about half of the total amount. Yttrium, neodymium, and lanthanum are next in order; samarium, gadolinium, praseodymium, dysprosium, erbium, holmium, and ytterbium are increasingly scarce; while the elements lutetium, thulium, terbium, and europium are extremely rare. The rare earths as a group are metals, usually trivalent, forming basic oxides, and insoluble oxalates in dilute mineral acids. Their fluorides are also difficultly soluble; hence the rare earths or lanthanides may be separated, in general, from other elements by adding oxalic or hydrofluoric acid to a dilute mineral acid solution containing them. The most striking fact which characterizes these elements is the remarkable .similarity in both the physical and chemical properties of their compounds. Their main differences are in the solubilities of their salts and the basicity of their oxides, which vary between that of the alkaline earths and that of aluminum.

Jan 1, 1960

By Clay Haynes, Tim Pearia, Robert Robinson, Judy Cochran

A 1.89 km-long, 4.5 m-I.D. tunnel was successfully constructed through saturated hard clay, and highly abrasive dense sand and gravel, with groundwater heads ranging from 10 to 20 m. In order to address these conditions, a Lovat earth pressure balance TBM was used, with soil conditioning by foam, water and various polymers. Support was provided by a single-pass segmented concrete lining. Groundwater inflows of 400to 1,600 lpm at times created challenges for changing cutting tools. Surface settlements were generally less than measurement thresholds, except under very low cover at startup. This project provides a pivotal case study of the successful application of EPBM technology in saturated, abrasive glacial soils and the successful application of soil conditioners.

Jan 1, 2003

By J. O. Voigt

No major changes took place in the salt industry during 2004. No new uses of salt appeared and only localized supply disruptions occurred. Many salt producers reported slightly increased salt sales and profits for the year. In the United States, about half of the salt supply is produced as brine, by solution mining, and is used by the chemicals industry. One-third of U.S. salt is produced as rock salt by 14 underground mines. The remaining two production techniques, each with less than 10 percent of the total tonnage, are vacuum-pan-evaporation for refined, food-grade salt, and solar pond crystallized salt.

Jan 1, 2005

By Sunny Schoone, Maximilian Getz, Felix Leaman, ELISABETH CLAUSEN

Jan 5, 2023

By Steven G. Vick

Effects of the August 1995 failure of the Omai tailings dam in Guyana continue to be felt throughout the mining industry. This article is taken from the final report of the independent body of experts commissioned by the government of Guyana to establish the cause of the failure. It explains what happened, why and how to avoid it in the future. Despite environmental, health and safety effects that were negligible by any objective measure, the failure has received wide international attention. Coming on the heels of the shift in exploration targets and development opportunities to developing countries, it has intensified interest in tailings dam safety and environmental management issues. An informed understanding of the nature of the failure and its effects is key to responding to these concerns.

Jan 1, 1996

By Chester A. Rowland, David M. Kjos

INTRODUCTION Mineral ore comminution is generally a feed preparation step for subsequent processing stages. Grinding, the fine product phase of comminution, requires a large capital investment and frequently is the area of maximum usage of power and wear resistant materials. Grinding is most frequently done in rotating drums utilizing loose grinding media, lifted by the rotation of the drum, to break the ores in various combinations of impact, attrition and abrasion to produce the specified product. Grinding media can be the ore itself (autogenous grinding - primary and secondary), natural or manufactured nonmetallic media (pebble milling) or manufactured metallic media - steel rods, steel or iron balls, or a combination of autogenous media and steel balls (SAG milling). This chapter covers rod and ball mills utilizing manufactured metallic grinding media. MILL DESIGN The interior surface of rod and ball mills exposed to pulp and/or grinding media are protected from wear and corrosion by rubber, metallic or a combination of rubber and metallic wear resistant materials. Rod and ball mills essentially draw constant power, thus are well suited for use of synchronous motors with power factor correction capabilities as drive motors. A net of approximately 120 to 130 percent of running torque is required to cascade the charge in these mills. The pull-in torque is about 130 to 140 percent with the pullout torque to keep the motor in-step (in-phase) generally in excess of 150 percent. When rod and ball mill are started across-the-line the starting and pull-in torques result in inrush currents exceeding 600 percent which possibly result in high voltage drops. To deliver 130 percent starting torque to the mill the motor design must take into account the maximum anticipated voltage drop. Motor torque decreases as the decimal fraction of the voltage available squared. E.g., a motor rated 160% starting torque with a 10% system voltage drop will deliver 160% x or 129.6% torque to its output shaft When it is not possible or practical to start a fully loaded synchronous motor across-the-line it is possible to utilize the motor's pull- out torque to start the mill. By using a clutch, normally an air clutch. between the motor and the mill, the motor is brought up to synchronous speed before the clutch is energized. If the motor has an adequate amount (175 or greater) of pull-out torque the pull-out torque starts the mill without major disruptions on the electrical system. Since the energy release at initial cascade of the mill charge is an inverse function of acceleration time, a minimum acceleration time of 6 to 10 seconds or more is recommended to prevent damage to the mill or the mill foundation. Economics at the time of plant design and mill purchase determine the drive to be used. The simpliest drive is the low speed synchronous motor with speeds in the range of 150 to 250 RPM connected to the mill pinionshaft by either an air clutch or flexible coupling. Using a speed reducer between the motor and pinionshaft permits using motors having speeds in the range of 600 to 1000 RPM. In this speed range, if power factor correction is not required induction motors can be used; squirrel cage where there is no restriction on inrush current; slip ring where a slow start and low inrush current is required. Air clutches can also be used to ease starting problems with squirrel cage motors. In some areas of the world induction motors and starters are less expensive than synchronous motors at a sacrifice of motor efficiency and power factor correction. Dual drives, that is two pinions driving one gear mounted on the mill, become economical for ball mills drawing more than 3500 to 4000 horsepower (2600 to 3000 kilowatts). Further developments of the low frequency, low speed synchronous motors with the rotor mounted on the mill shell or an extension of one of the mill trunnions could improve the cost picture for these "gearless drives", making them practical for large ball mills. The percent of critical speed, which is the speed at which the centrifugal force is sufficiently large to cause a small particle to ad- here to the shell liners for the full revolution of the mill is given in mill specifications. Critical speed is determined from the following: Where D is mill diameter inside liners (specified in meters). Cs is critical speed in RPM. When D is specified in feet

Jan 1, 1998

By Stephen M. Smading

The Jinping-II hydroelectric project in China features four parallel headrace tunnels approximately 18-km- (11-mile-) long. Two will be excavated by tunnel boring machines (TBM) and two by drill-and-blast. A nearby fifth tunnel is being excavated by a 7.2-m (23.6-ft) TBM to draw down ground water in advance of excavating the headrace tunnels. Unique onsite assembly of a 12.43-m (40.7-ft) main beam TBM and backup system was completed on Sept. 18, 2008 in the remote mountains of the Sichuan province. The equipment was assembled onsite, without previously having been assembled and tested in a factory, using a method called onsite first time assembly (OFTA). This paper addresses the following topics: ? Project description including geology and terrain. ? The decision process leading to onsite, first time assembly, the assembly process and logistics, and lessons learned. ? Design features of the TBM and backup system. ? Operational history from startup to the writing of this article.

Jan 1, 2009

By S. Acar, M. Botz

Introduction In 2007, Newmont Mining Corporation will begin construction of a milling and cyanidation leaching circuit for the Yanacocha site in Peru. Ore to be processed in the circuit contains a relatively high level of cyanide-soluble copper, and as a result, leach solutions are expected to contain elevated levels of copper and WAD cyanide. As the means of controlling copper levels in leach solution, a circuit will be installed to precipitate copper sulfide from barren solution using sodium hydrosulfide. In addition, a circuit will be installed to recover cyanide from barren solution, with recovered cyanide to be returned to the leaching circuit and treated solution to be utilized as make-up in existing heap leaching facilities. As needed to support design of the copper precipitation and cyanide recovery circuits, pilot testing was conducted in 2005 to determine preferred operating conditions and evaluate the performance of the processes. This paper presents results from the pilot testing where 99% copper reduction was achieved and 95% to 98% cyanide recovery was achieved.

Jan 1, 2007

By Jeffrey A. Pontius

The Cresson gold deposit is in the southcentral portion of the Cripple Creek mining district in central Colorado (Fig. 1). The deposit is hosted by a series of 32 Ma (Kelley, 1995) felsic to mafic, alkaline volcanic rocks that were deposited in a diatremal environment. Mineralization in the Cresson deposit is structurally con-trolled. It occurs as disseminated and vein type. Gold deposition is closely associated with potassium-pyrite metasomatism and quartz-telluride vein development(Pontius, 1992). The Cresson deposit is owned by the Cripple Creek and Victor Gold Mining Co. (CC&V), a joint venture between Pikes Peak Mining (67%) and Golden CycleGold (33%).

Jan 1, 1996

By Ali Sogue, Steven Schurman, R. Jensen

This paper presents the results of the remediation of the Gregory Gulch #1 Superfund site, an EPA Clear Creek Superfund project located in central Colorado. The Gregory Gulch #1 site was an historic gold mine and mill located within the city limits of Central City, Colorado. Remediation efforts included the use of the Envirobond™ additive to chemically bind the metals in 3,000 tons of acid generating mine waste that contained high levels of arsenic, lead, zinc and mercury. The toxicity of the mine waste amended with Envirobond™ was not only reduced to RCRA TCLP standards but also to the more stringent UTS standards. The final cost of the treatment was much less than standard industry solidification processes and significantly less than offsite disposal options.

Jan 1, 1999

By Dan Mackie

Commissioning of a 150 tpd continuous vat leaching pilot plant at Ariel Resource's San Martin Mine began in October, 1995. Functioning of the systems proved to be as designed. Initial tri­als, however, were besieged with problems associated with clay fractions that were much higher than in the ore originally tested. Nevertheless, leach rates were good in the initial and subsequent trials. Difficulties with the viscous solutions generated by the clay fraction continued, requiring modifications to the solution collection methods and investigation of flocculant types and mixing points. Trials with deslimed ore and sand feed demon­strated continuous flow. Current modifications will be tried in late May, 1996, along with selected flocculants, with the objec­tive of leaching continuously without the need to deslime.

Jan 1, 1996

By Timothy P. Smirnoff, Anthony Stirbys, Amanda Elioff

The construction of the continuing expansion of the Los Angeles Metro System has required the construction of eight new station structures. One contractor proposed, as a construction alternative for the support of excavation of the Hollywood/Vine Station, the use of soil-mix walls instead of the conventional soldier piles and wood lagging. Soldier piles and shotcrete or wood lagging support systems had been generally used on the other sub-way station excavations constructed to date throughout the system. Immediately east of that station and in almost exactly the same geology, the Hollywood/Western Station was being excavated at the same time using a conventional support system. This paper com-pares the performance of these two support systems in the alluvial soils of the Los Angeles Basin.

Jan 1, 1997

By R. D. Hendry, R. A. Pearce, Allyn J. Steward, Dennis C. Martin

This paper outlines the geotechnical exploration and design approach utilized for the Km 80 and Km 86 Tunnels on the British Columbia Railway Tumbler Ridge Branchline. Information from surface mapping, trenching and diamond drilling was used to prepare a rational support design using accepted rock mass classification systems. Four possible support types were developed and used as a basis for preparing a cost estimate and contract documents for the tunnels. A unit price contract was selected to provide flexibility and to enable rational selection of support based on actual ground conditions encountered during excavation. Although ground conditions along the tunnels are variable, the general geology and quantities of support installed were consistent with those estimated from the geotechnical investigation and design.

Jan 1, 1985

By E. B. McNew, J. M. Barnes, A. E. Torma

The applicability of biosorbent materials for the removal of uranium and other heavy metals from synthetic and actual uranium mil 1 tailing (UMT) solutions has been investigated. The biosorbent materials used in this study were barium loaded beads of alginate, alginate-dextran sulfate, alginate-carrageenan, and alginate-xanthan gum. These studies were performed in stirred vessels and packed columns. It was observed that 100 percent of the uranium could be removed from a synthetic solution following 2 hours of contact time in a stirred vessel, and 80-100% of U, V, Mo, Mn, Sr, and Zn could be removed from a UMT effluent solution with a packed column. Kinetic evaluations were performed, using a generalized rate equation, on the biosorption of uranium with these barium loaded beads.

Jan 1, 1992

By H. Cho, P. T. Luckie

A study was undertaken to investigate the liberation behavior of coal macerals through size reduction and density separation. It was seen that the liberation of coal macerals could be achieved by size reduction, allowing selective separation of coal organic components. However, the degree of liberation was dependent upon the comminution history. Even for the same fineness, the extent of liberation might not be the same depending on how the material was generated. The liberation of coal macerals was analyzed by examining the liberation index. It was seen that, although comminution produced fine particles enriched in inertinite (primarily fusinite and semifusinite), these particles were less liberated than coarser particles. However, the liberation of macerals increased when all the feed was ground to a finer size.

Jan 1, 1999

By P. James, L. J. Steiner, F. Turin

NIOSH researchers and an industry partner conducted an investigation to study the effectiveness of ergonomic interventions in mining. The study required a significant commitment of time and resources from the cooperating site, Bridger Coal Company. This paper focuses on some of the challenges faced by the researchers and their cooperator. Challenges included the need to redesign methods of identifying and prioritizing injuries, and the need to train the workforce in ergonomics while still maintaining progress and interest in the process. The authors discuss some practical interventions implemented by Bridger Coal to date, and discuss future research implications brought about by PacifiCorp’s Generation Business Unit, who monitored the project to determine whether an employee-owned ergonomic process could be implemented throughout their operations.

Jan 1, 2003

By Charles A. Maliszewski, James M. Galvin, Gerhard Urschitz

With obvious signs of deterioration, the double track, historically registered Bergen Tunnels as part of the commuter rail system in New Jersey waited for rehabilitation for a long time. Serious water in filtration not only caused operational and safety hazards for the electrified rail line but also deteriorated and structurally compromised the tunnel lining. An innovative rehabilitation program was developed to enlarge the tunnel and to provide a new waterproofing system and final lining. For the rock excavation and support New Austrian Tunneling Method (NATM) principles were utilized, while water tightness was achieved by installing a PVC membrane waterproofing system followed by a reinforced cast-in-place final lining. Experienced on-site supervision generated additional savings during construction by streamlining support elements based on geologic conditions encountered.

Jan 1, 2003

By George Weinman

GEORGE WEINMAN: Our first group leader for the discussion, "Ups and Downs of Downsizing," wishes to remain anonymous. ANONYMOUS PRESENTER: When George and Jamie asked me to do this, I thought a while about what to say and went through several variations of what I’ll present today. I came up with the title "Ups and Downs of Downsizing." A lot of downsizing has occurred in all of our industries-mining in particular. While this downsizing has gone on there's no doubt that all of us produced more and more, probably with less and less people and less and less resources. But I contend that there's also a big downside associated with this. I disagree with a point made in the keynote presentation yesterday, the point being that vendors are bringing new technology to us. I can agree that the major changes associated with equipment and similar items are being brought to us by vendors. But I do contend that there's an awful lot that's getting away from us-especially if you factor in the integrated supply or the strategic alliance or the various and sundry related purchasing techniques. I contend that we are missing a lot of, let's call it, lesser technology; not major equipment changes, but different tools, different techniques for doing things, different left-handed widgets. The integrated suppliers, to some degree, are bringing this to us; but I am concerned about the small guy who comes up with a good idea but deals through integrated suppliers rather than having direct contact with plant people. They're probably going to go somewhere else and market their wares. That's what I'm hearing from a lot of smaller vendors. I also think that we are losing a lot of our knowledge base. Every time you shrink an organization somebody leaves. The bottom line: less people. When a person leaves-and I don't care who he or she is, be it the janitor who's cleaning the bathroom, the operator, the engineer-when that person leaves there's some knowledge that goes with that per- son. And sometimes that knowledge is quite valuable. I've worked more or less at the plant "low." I can't tell you how many excellent, viable projects have come from people who said, "Look here. There's a better way of doing this." And, you know, they were right. Now, maybe they didn't have the resources or the expertise to develop these ideas into full projects, but they had that basic idea. Someone may say, "Well, if we put a no. 6 skyhook here, so we can lift this particular pump without bringing a crane from 30 miles away ...." No. 6 skyhooks are a bit expensive, but what if we put together a little structural steel here or whatnot?

Jan 1, 1998

By Paul D. Dupree

Introduction Trapper mine is a dip-line strip coal mine, located about 10 km (6 miles) southwest of Craig, CO. The mine provides the Craig Generating Station, operated by Colorado-Ute, with about 2 Mt (2.3 million st) of coal annually. The Trapper mine is situated on a hillside that climbs in elevation from 2 to 2.5 km (6500 to 7400 ft) at average grade of 15%. The coal seams dip at about 16% down-grade and they outcrop near the top of the hillside. They exceed maximum stripping depth in 1 to 1.7 km (3500 to 5500 ft) of pit length. Stripping in the three existing pits is performed with three Page 752 23 m3 (30 cu yd) walking draglines. Two of the pits are currently being cast blasted. Until the active cut, the Derringer pit was cast blasted only where the lower mined seam exceeded stripping capabilities and the pit was a single seam operation. Now, a 9-m (30-ft) interburden bench has been incorporated into the stripping plan. This bench protects the first seam of coal from shear when the overburden is cast blasted. The cast blasted section of the 1.6 km (53,000 ft) long pit has been increased from the 457 m (1500 ft) casted previ¬ously to 1 km (3500 ft). High speed films show the coal losses due to shear of the blast are minimal. The pit width in Derringer pit is 61 m (200 ft). This is necessary for efficient spoil side stripping of a relatively thin interburden in the two-seam section of the pit. The Enfield pit is entirely cast blasted. The single seam pit is 46 m (150 ft) wide, 1.1 km (3600 ft) long, and varies from 15 and 46 m (50 and 150 ft) in overburden depth. The overburden strata at the mine consists of medium-hard layers of sandstone, shale, and

Jan 1, 1987

By J. A. Finch

The photographic measurement of bubble diameter is tedious and restricted to vessels with transparent walls and relatively low bubble concentrations. A new method for predicting bubble diameter in a flotation column has been developed and confirmed by comparison with measured mean bubble diameter (Yianatos et al., 1987) Flotation columns are not mechanically agitated. This note describes an extension of Ms method to a mechanically agitated flotation machine. A recent publication by Szatkowski (1987) provided sufficient data to calculate bubble diameter by the proposed method and compare with values determined photo- graphically. Method In a two-phase system (gas and liquid), bubble diameter (db), superficial gas rate (Vg) and liquid rate (Vl), and gas holdup (Eg) are linked by two relationships (see the nomenclature at the end of the paper for definitions of symbols). [ ] Equation 1 is the definition of slip velocity (relative gas to liquid velocity) for a countercurrent gas-liquid system, and equation 2 is an adaptation (Yianatos et al., 1987) of the expression of Masliyah (1979) for hindered settling in a multispecies system. The solution for bubble diameter is by repetitive substitution of estimates of db in equation 2 until the calculated Ubs is equal to the measured Ubs in equation 1. Results and discussion Table 1 presents the data obtained by Szatkowski (1987) from a mechanically agitated laboratory flotation machine. The machine generated bubbles by directing a gas stream into the vicinity of a disk rotating at Ngh speed (up to 8000 rpm) in the flotation slurry. The air was completely dispersed in a swarm of fine bubbles. A photographic technique was used to determine the distribution of bubble size. The mean diameter and standard deviation were then calculated. Table 1 provides all the information necessary for computing bubble diameter (Vl is assumed zero). The calculated bubble diameter is presented in Table 1 as d;. Figure 1 plots measured versus calculated bubble diameter and shows reasonable agreement between values. There is an apparent tendency to undercalculate the small bubbles and overcalculate the large ones.

Jan 1, 1987