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By Jay J. Burns

TWO zinc concentrating mills are operated by the St. Joseph Lead Co. in St. Lawrence County, N. Y. The Edwards mill is operating at present only sixteen hours a day treating 400 tons daily. The metallurgy at Edwards presents no unusual problems therefore this discussion is mainly limited to the metallurgy of the Balmat ore. The Balmat mill has a capacity of 1,250 tons per day and the operating crew consists of three shift bosses each with five millmen per shift. Two crusher operators work on each of two shifts and one cleanup man on day shift. The repair crew is made up of a foreman and three repairmen, one of the repairmen being capable of taking over any operating job in the mill from shift boss down. Each millman is capable of handling the next job in line and several men in the yard crew are broken into handling the start-off job in the mill so that temporary shortages of men do not cause extensive overtime. Sphalerite, pyrite, and galena, the principal ore minerals, occur in a gangue of calcite, dolomite, and silicates, especially diopside, plus tremo-

Jan 1, 1947

By M. Dupuis

Recently, a new approach to controlling alumina feeding in a H.H. cell has been proposed [1, 2 and 3]. This approach, labeled In Situ aluminium cell control, is based on a correlation between the slope of the cell voltage during a no feed observation period and the dissolved alumina concentration in the bath. Once the dissolved alumina concentration has been estimated, a primary calibration surface uniquely relating the dissolved alumina concentration and the anode-cathode-distance (ACD) to the cell voltage is used to estimate the cell ACD. Once both the alumina concentration and the ACD have been estimated at the end of the observation period, a PID controller can be used to regulate the rate of alumina feeding using only the normalized cell voltage to calculate the error between the estimated dissolved alumina concentration and a target value. The In Situ aluminium cell control algorithm has now been tested with the Dyna/Marc [4, 5 and 6] cell simulator demonstrating that it should be possible to control a H.H. cell in this manner with less variation in the dissolved alumina concentration and other process variables producing a prediction of improved cell current efficiency and power efficiency.

Jan 1, 2011

By Yuhua Wang, Haitao Zheng, Guangli Zhu, Yuehao Zhao, Fushun Yu, Dongfang Lu

"Spodumene, feldspar and quartz possess similar surface properties, which make the difficult to separate spodumene from other silicate minerals by flotation. The differences of crystal structure, surface property and floatability of three minerals were studied by various technical means. The results show that Li-O and Al-O bonds in spodumene structure are not strong and easy to be broken. However, Al-O and Si-O bonds in feldspar and Si-O bond in quartz are very strong and not easy to be broken. High contents of Al and Li on spodumene surface were found by XPS measurements, which is benefit for the adsorption of anionic collector. The surface energies of main cleavage planes of spodumene (110), feldspar (001) and quartz (001) are 1.88 J/m2, 2.97J/m2 and 2.68J/m2, respectively. Hence, the exposed surfaces of feldspar and quartz are not saturated and easy to adsorb OH- ions from solution and cause the lower IEP than spodumene. The contact angles of three minerals in water are lower than 16º, suggesting that the hydrophilicity of three minerals is high, and the order of hydrophilicity is quartz > feldspar >spodumene. Correspondingly, contact angles and flotation recoveries of spodumene are bigger than that of feldspar and quartz in presence of a combined anionic and cationic collector named YOA.INTRODUCTION Flotation technology plays most important role for the recovery of spodumene from pegmatite ore. Generally, spodumene with fresh surface is easy to be floated using soap and oleic acid as collector, and the best flotation pH range is 4.0 to 9.0. However, the flotability of spodumene will become worse when its surface was polluted by slime which produced by the weathering of spodumene and other silicates (Dresler et al. 1998; Moon and Fuerstenau 2003; Menéndez et al. 2004; Yu et al. 2015). In addition, the species of Ca2+, Fe3+, Mg2+ and other metal ions in pulp can not only activate the flotation of spodumene, but also the activation of gangue minerals, and then the flotation selectivity of spodumene from gangue minerals became wores (Wang and Yu 2007; Yu et al. 2014). In order to eliminate the negative effects of slime and metal ions on the flotation of spoddumene, the desliming and long time scrubbing with the addition of NaOH and Na2CO3 were adopted. Usually, NaOH shows the activation for spodumene flotation, and the recovery of spodumene increases with the increase of NaOH dosage. Na2CO3 can eliminate the non selective activation of Ca2+, Fe3+, Mg2+ ions in the pulp, and the concentrate grade of spodumene increases with the increase of Na2CO3 dosage. Anionic collectors such as chelants, hydroxamic acid, oxidized paraffin soap, naphthenic acid and oleic acid were widely used for the flotation of spodumene (Dresler et al. 1998; Moon and Fuerstenau 2003; Menéndez et al. 2004; Yu et al. 2015)."

Jan 1, 2018

By In-Hyeok Choi

Currently, spent V2O5–WO3/TiO2 catalysts contribute significant amounts of solid waste following increasing levels of global demand. V2O5–WO3/ TiO2 catalysts usually consist of TiO2 anatase as a supporting oxide, vanadium as a catalytic agent, and other promoters such as tungsten, silicon, and calcium. Although vanadium is the main catalytic agent, a relatively high content of tungsten (WO3, 7–10 wt%) typically exists compared to vanadium (0.5–1.5 wt%). Considering the irreplaceable properties and industrial importance of tungsten, a feasible method for the recycling of spent V2O5–WO3/TiO2 catalyst should be established to utilize it as a secondary source. This paper presents a process to recover tungsten from spent V2O5–WO3/TiO2 catalyst. The processes proposed here involving roasting, decomposition using HCl, ammonia leaching, and crystallization. Within the process, ammonium paratungstate (71 wt% as WO3) is obtained as a final product. The total yield rate of tungsten from feedstock was found to be 96.3%.

By I D. Gregson, A Gandhe, A S. Garcia, K M. MacAndrew

"Falls of ground still cause around one-third of all underground mining fatalities. The recent introduction of mechanised room and pillar working into Indian coalmines is an example of the application of new design by measurement and safety monitoring strategies, which can be used to reduce falls of ground risks in mines.Rock Mechanics Technology, now part of Golder Associates UK Ltd (Golder), is providing rock bolt support and pillar design services to the mines using a design by measurement approach. This involves in situ monitoring to confirm support and pillar designs and pillar extraction sequences.Site examples are given of the support designs determined and the monitoring programs undertaken for both design and safety.Routine safety monitoring of roof movement using telltales is also described. A new auto warning telltale has been developed specifically to warn of imminent caving as pillars are extracted.The approach used by Golder has played a major role in ensuring the success of these projects. As a consequence, mechanised room and pillar working is now firmly established in India, and wider application is planned to help meet future demand for coal.The experience confirms that the design by measurement method for this application has important advantages over conventional design approaches using analytical or empirical methods. Firstly, it measures the design performance or outcome rather than trying to calculate or predict it. Secondly, it provides the opportunity for the design to be adjusted to obtain the optimum outcome. Thirdly, it can do this at a local level, so that the support design can be matched to changing conditions to give the most cost-effective design. Finally, it incorporates long-term monitoring of support performance, again at a local level, so that subsequent change can be identified to enable corrective measures to be implemented in a timely manner and to ensure that ground control safety is maintained.CITATION:MacAndrew, K M, Garcia, A S, Gregson, I D and Gandhe, A, 2014. Application of design by measurement for support and pillar design in mechanised room and pillar coal mining in India, in Proceedings AusRock 2014: Third Australasian Ground Control in Mining Conference, pp 65–76 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Nov 5, 2014

Jan 1, 1968

Flotation testing has been done at laboratory scale on hematite mineral and ore samples using stage additions of various hydroxamates. Chemisorption of hydroxamate species onto the hematite surface was inferred from the results, possibly involving both the undissociated hydroxamic acid and hydroxamate ion as maximum recovery occurred close to the pKa of the hydroxamic acid. Flotation testing on a specular hematite ore containing 40 per cent Fe and 42.5 per cent silica gave highest iron grades in concentrate at pH 6, in contrast to published data which suggested pH 9 as the optimum value for ore processing. Using octyl hydroxamate, a rougher concentrate assaying 67 per cent Fe at 51 per cent recovery was obtained using four stage additions of 50 g/t at pH 6. For lower grade concentrates still meeting the chemical analysis for pigment grade micaceous hematite (minimum Fe assay 60 per cent) iron recoveries of 80 per cent were possible for stage additions of 50 g/t octyl hydroxamate in the pH range 6-8.

Jan 1, 1999

By James H. Cobbs

The utilization of large diameter drilled holes for mine shafts seems to have had its beginning in Europe about 1850 (28) where the mining industries of Germany and The Netherlands first used drilled shafts. Recent shafts in softer formations of diameters as large as 25 feet and as deep as 1, 700 feet have been drilled in both The Netherlands and Germany. The first shaft drillers utilized cable tool or churn drilling techniques which limit both hole diameters and drilling speeds. Though some churn drilling of shafts is still performed, rotary drilling has been the dominant technique for many years.

Jan 1, 1973

Jan 1, 1970

The first of this three part series examines he history and some of the properties and characteristics of explosives. Use of explosives is discussed for underground and surface coal operations—the largest consumers in the mining industry. Part two of the series will deal with explosives use in surface and underground hardrock operations. Part three will discuss explosives transportation, storage, handling, and destruction.

Jan 7, 1981

By Wenbo Lu

A combined method is present in this paper in analysis the dynamic stability of rock slope under the loading of blast induced seismic, in which, by combining the dynamic load calculating method of quasi-static method with the load inputting method of dynamic finite element method, such factors as the decay of seismic wave along its way of propagation, the frequency. amplitude. the phase difference of seismic wave, etc., can all be taken into account in analysis of the dynamic stability of slope.

Jan 1, 1999

By AIME

Burton J. Westman-Besides decreasing the diamond size, there appear to be two other approaches open to overcome excessive diamond loss and, more particularly, the rapid diamond polish that took place in this quartz-feldspar rock. First, however, a short discussion of the diamond characteristics should be made in order to discuss the situation more clearly. The West African diamonds which we can assume were used. for this drilling have two disadvantages when drilling hard, exceptionally fine-grained rocks. These are rapid polishing and, if .the bit isn't "drilled in" carefully, a tendency to damage, resulting in excessive diamond loss. The normal wear resistance of any diamond is inversely proportional to the compressive strength of the diamond, therefore only the highest grade, of diamonds free from insipient fractures should be employed. Using upwards of 90 diamonds per carat instead of the20 to 30 per carat not only decreases the total projected area of the cutting points but increases the stress per point which, in turn, causes greater depth of penetration into the rock. Considering this depth of penetration, therefore, suggests an approach other than increasing the, number of diamonds per carat and that is to decrease the actual number of diamond points (also the carat weight per bit) so that the penetrating force on an individual diamond is increased. This penetration increases the chipping or rock shearing action.

Jan 1, 1950

By N Kawashima, R St C Smart, A M. Robertson

This paper focuses on a case study at a base metal mine located in a tropical semi-arid area of Western Australia (WA). It is demonstrated that understanding the geochemical nature of the mine tailings forms an integral part of innovative management strategies developed throughout the life-of-mine (LOM) when planning for mine closure. The tailings are currently stored at an above-ground valley fill tailings storage facility (TSF) and one option for long-term management under consideration is leaving the tailings in situ at mine closure and covering using the concept of an enhanced æstore and releaseÆ cover system.Extensive geochemical and mineralogical studies have found that the geochemical behaviour of the tailings in the TSF is significantly more benign than initially predicted. The sulfide content of the tailings is primarily pyrrhotite, which can be acid-forming, acid-neutral or acid-consuming depending on a number of factors including redox conditions. Tailings pore water at depth is pH neutral with excess alkalinity, low concentrations of dissolved metals and elevated sulfate concentration.Recent geochemical tests on tailings samples indicate that a reduction in the acid potential and a corresponding increase in the neutralising potential of the tailings, through partial removal of pyrrhotite at the ore processing stage, has the potential to further reduce the long-term risk of potential impact to the environment from any TSF seepage. The current work program investigates whether partial removal of pyrrhotite is feasible and secondly, whether this approach is necessary to produce a significantly better outcome at mine closure.CITATION:Robertson, A M, Kawashima, N, Smart, R St C and Schumann, R, 2012. Understanding tailings geochemistry - Continuous learning and innovation when planning for mine closure, in Proceedings Life-of-Mine 2012 , pp 287-292 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 10, 2012

Isometric strength is defined as the capacity to produce force or torque with a voluntary isometric (muscle[s] maintain[s] a constant length) contraction. The key thing to understand about this type of contraction and strength measurement is that no body movement occurs during the measurement period. The tested person's body angles and posture remain the same throughout the test. Isometric strength has historically been the type most studied and measured. It is probably the easiest to measure and to understand. Some strength researchers feel that isometric strength data may be difficult to apply to some "real life" situations because in most real circumstances people are moving - they are not static. Other researchers counter that it is equally difficult to determine the speed of movement of a person or group of persons doing a job (each moves in his or her unique manner and speed across the links and joints of the body). Thus, dynamic strength test data collected on persons moving at a different speed and/or in a different posture from the "real world" condition will be just as hard to apply. In truth, neither is better - they are different measurements. Both researchers and users should collect and use data that they understand and that fits their application. Workplace Assessment When a worker is called on to perform a physically demanding lifting task, the external load produces moments - tendencies to produce motion, also called torques - about various joints of the body.(1) Often these moments are augmented by the force of gravity acting on the mass of various body segments. For example, in a biceps curl exercise, the moment produced by the forearm flexors must counteract the moment of the weight held in the hands as well as the moment caused by gravity acting on the forearm's center of mass. To perform the task successfully, the muscles responsible for moving the joint must develop a greater moment than the combined moments of the external load and body segment. It should be clear that, at each joint of the body, there is a limit to the strength that the muscle can produce to move ever-increasing external loads. This concept forms the basis of isometric muscle strength prediction modeling.(1) The following procedures are generally used with this biomechanical analysis technique. First, workers are observed (and usually photographed or video-taped) as they perform physically demanding tasks. For each task the posture of the torso and the extremities are documented at the time of peak exertion. The postures are then re-created using a computerized software package, which calculates the load moments produced at various joints of the body as the task is performed. The values obtained during this analysis are then compared to population norms for isometric strength obtained from a population of industrial workers. In this manner, the model estimates the proportion of the population

Jan 1, 1998

By T. M. Garland

This report presents a more extensive perspective of the calculated profitability of developing and producing crude oil and associated natural gas by primary methods from selected onshore areas of the United States. Estimates were made for cost of drilling, completing, equipping, and producing oil wells at four depths in seven onshore producing areas. A computer program was developed for calculating the discounted cash flow rate of return for the 64 different models in each area resulting from varying the oil-producing rate and production decline rate. The results of these calculations provide data that vividly illustrate the change in rate of return to changes in well depth, initial producing rate, and production decline rate, as well as to variations in development cost and operating expense. Cost per barrel of oil produced in terms of annual cost and total cost is also discussed. Included in the report are explanations of the method of determining costs and the procedure used for profitability analysis. Summaries of all cost estimates and results of profitability analyses are presented in the appendix in tabular form for each of the seven producing areas.

Jan 1, 1972

By Mike Wongkaew, Brandon Kluzniak, Per Vilhelmsson

Rail access to Ontario International Airport (ONT) has long been a goal of San Bernardino County Transportation Authority (SBCTA) to increase public transit options to ONT and provide incoming ONT passengers improved access to San Bernardino County and the greater Southern California region. This paper describes the planning for the Design-Build-Operate-Maintain (DBOM) delivery of an emerging technology tunnel connection between the Rancho Cucamonga Metrolink station and ONT. The paper discusses key technical considerations, challenges, and opportunities; conceptual configuration development; and construction planning and establishment of the footprint for the underground connection.

Dec 1, 2022

By Peter K. Strangway

During 1973, a 20,000 net ton (18 100 metric ton) formcoke test was carried out at Inland's 26.5-foot (8.08-meter) hearth diameter on NO. 5 Blast Furnace. The formcoke briquettes were produced from Elko1 coal by FMC at their pilot plant in Kemmerer, Wyoming. The briquettes were shipped to Inland in open railroad cars and most of them were placed in an outside storage field until a sufficient amount was available for the actual blast furnace test. The test consisted of three phases: an initial four-week base period; the actual 33-day formcoke test period; and a final three-week post-test base period. During the formcoke test period, the regular coke skip loads were gradually replaced with skips of formcoke until, after a three-week period, the replacement ratio reached 80%. The furnace operated smoothly and quality hot metal was produced. The burden remained permeable and it was possible to maintain wind rates 9 of over 100,000 scfm (47 Nm3/s) without significant blast pressure increases. Dusting from the formcoke was not a problem either in the stockhouse or in the top gas. The formcoke briquettes maintained their original shape and density as they descended through the furnace. When the 80 percent replacement level had been reached, the inwall temperatures became higher than normal (about 2000 ºF (1090 ºC)) and followed an unusual steady pattern. In addition, the top gas temperature became higher than normal, and the hot metal analyses and temperatures began to fluctuate unacceptably. In general, it is felt that the high inwall temperatures were a result of channeling of the hot furnace gases along the stack walls. This channeling appears to have been promoted by a change in distribution of the burden in the top of the furnace which, undoubtedly, was caused by the increased increments of formcoke. Since it was felt to be unsafe to operate the furnace for an extended period of time with the inwall temperatures at the high level, a decision was made to remove the formcoke from the furnace. When this was done, the inwall temperatures returned to normal. Once the inwall temperatures had returned to normal, the remaining 3,800 net tons (3450 metric tons) of formcoke were used in the furnace at a nominal 40% replacement level for five continuous days. The furnace operated smoothly during this period, and the results were representative of satisfactory furnace operation with partial replacement of regular coke with formcoke. The test was successful in proving that formcoke can replace up to about 50% of the regular coke in a large furnace operating at high wind rates. However, additional work will be required to develop acceptable charging practices for using higher levels of formcoke replacement.

Jan 1, 1977

By C. Vanvoren

"Since 1975, the Aluminium Pechiney Research Centre at Saint-Jean-de-Maurienne (LRF) has been concentrating its efforts on designing, testing and perfecting the Aluminium Pechiney point-fed prebaked reduction cell.In 1979, the AP18 cell (180 kA) was industrially launched with the construction and start-up of the F line in Saint-Jean. The following generation, the AP30 (300 kA), with prototypes started in 1981, was launched in 1986 on the Saint-Jean G line. Along with cell design, efficient process control systems were developed to pilot alumina feeding and bath composition. Together, these pots are currently producing 4.6 Mt of aluminum, i.e., 18% of the world’s overall yearly production.In 1989, the research centre faced a new challenge: to push the limits for the size of a reduction cell to further improve the profitability of Greenfield projects. This new generation development was also the opportunity for a significant redesign of the Aluminium Pechiney electrolysis information system.With a track record of ten years of semiindustrial operation with three prototypes, the AP50 technology is now ready for industrial implementation and enables a significant breakthrough to be made in terms of return on investment of a Greenfield smelter."

Jan 1, 2002

By René Fernández

Newmont Mining Corporation has been processing refractory gold ore by roasting since January 1995. The ore is refractory due to the association of the gold with iron sulfides, namely pyrite and marcasite, and the presence of carbonaceous components and is oxidized in an oxygen-enriched gaseous atmosphere. The structural characteristics of the calcine have a strong influence on the eventual recovery of gold by cyanidation. These characteristics are strongly dependent upon the roasting conditions, mainly the roasting temperature.

Aug 25, 2002

By A. L. Crandell

History The Cortez Mining District has a long history of mining activity dating back to 1862. More recently, Cortez Gold Mines had operated a cyanide leach, counter current decantation (CCD) recovery gold mill from January 1969 to February 1976. This type of plant has been standard in the US for treating higher grade, finely disseminated gold in Carlin type ores. The mill was initially operated from 1969 to 1973 at 1.5 kt/d (1650 stpd) on oxidized ore from the Cortez ore body. In 1973, the mill was operated at 2 kt/d (2200 stpd) on softer, less oxidized, high clay ore from Gold Acres, located 13 km (8 miles) west of Cortez, NV. In February 1976, the oxidized portion of the ore body at Gold Acres was depleted and the mill ceased operations. Successful heap leaching of the lower grade portion of the ores was practiced at both Cortez and Gold Acres to supplement the milling operations. Valuable experience was gained in the operation of these two heap leaching programs (Duncan and Smolik, 1977). The ore heaps at Gold Acres and Cortez were releached with a limited crew from 1976 to 1979. The surge in the price of gold in late 1979 sparked an interest to incorporate new techniques into treating lower grade and less oxidized carbonaceous ores at Cortez. In early 1980, heap leaching was started on low grade oxidized dump material left from the earlier Cortez mining operation. By June of 1980, testwork was completed on the London Extension Dump, a low grade, mildly carbonaceous dump at Gold Acres. It was decided to convert the mill at Cortez from the conventional cyanide leach CCD recovery plant to a carbon-in-leach (CIL) recovery system. In May 1983, ore processing from the Horse Canyon deposit began. The Horse Canyon ore body contains both oxide and carbonaceous ores and is located 22.5 km (14 miles) from the Cortez mill. The Cortez CIL circuit was expanded to handle a blend of oxide and carbonaceous ores from Horse Canyon. Ore selection In the last quarter 1979, gold prices were increasing to a point where restarting the Cortez mill, idled since 1976, appeared imminent. Several previously explored geologic reserves at various stages of development were available as possible sources of feed for the mill. Initial metallurgical testing of dumps and tailings from earlier operations was completed, and previously uneconomic deposits were explored. It was decided to pursue sources of ore with the least lead time to production and those with the least capital cost requirements. A master plan was developed to bring the four profitable ore resources into production. The plan called for: • Heap leaching of waste dumps from the Cortez mining operation; • Restarting the mill to handle the waste dumps from the London

Jan 9, 1986