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By J. Elbrond

"This paper describes qualitatively the progressive are losses and rock dilution in a mining project from the geological modelling to and including the concentration. A simple theoretical model of the net present value of a deposit, assuming a log-normal grade frequency function, is used to find the pair of decisions on rate of production and cut-off grade, which maximizes the net present value. Various degrees of ore losses and rock dilution are then applied on this optimum, and the resulting decreasing net present values are calculated, as well as the increasing cost of metal and the increasing metal losses as a function of are losses and rock dilution.IntroductionDue to a variety of uncertainties, to inevitable lack of precision in deposit estimation and in mine planning, and due to production constraints, ore losses and rock dilution occur throughout the many phases of a mining project. The phases range from the geological definition of the orebody, over the cut-off grade optimization and decision, over the design of the mining method and the actual mining to the ore dressing, where the final loss of valuable ore into the tailings occurs. Quantification of ore losses and rock dilution are mostly difficult if not simply impossible to obtain. At the end-phase an estimate of the total dilution may be obtained with some degree of precision by observations of the amounts and analysis of concentrate and tailings. It is much more difficult to estimate the total ore losses because the uncertainty of the geological modelling in the initial phase and in the mining phase is very difficult to quantify, despite the use of geostatistics for grade estimation.Many attempts at quantification of ore losses and rock dilution have been made and some examples are shown in Table I. Although they differ widely from each other, it is obvious from these figures that ore losses and rock dilution are significant and have considerable effects on the economical results of a mining operation. Life length, net present value, cost of producing metal and the loss of metal are all affected."

Jan 1, 1994

By Yuhua Wang, Wei Sun, Dongfang Lu, Yuehua Hu, Tao Jiang

"With the continuous depletion of high-grade tungsten ores, the low-grade refractory wolframite has become an important source for the tungsten industry. However, the large-scale utilization of low-grade refractory wolframite is disappointedly scarce, it is due to leanness of valuable minerals, complex mineral characterization and use of toxic collectors.The aim of this study was to develop an environment-friendly processing flowsheet, which was used for beneficiation of low-grade tungsten ore containing 0.45% WO3 from New Brunswick in Canada, and the content of arsenic is 0.68% in this materials. The objective of beneficiation studied was to obtain a concentration containing 60% WO3 with less than 0.5% of arsenic under the condition without any flotation reagents. In this environment-friendly processing flowsheet, magnetic separation technology, classification and gravity separation technology were combined and employed. Firstly, magnetic separation was employed for the first pre-concentration process, over 95% of the feed was rejected as the final tailings, meanwhile, over 85% of arsenic minerals were removed with tailings. Secondly, gravity separation tests were employed on the classified magnetic pre-concentrate in the second pre-concentration process, a second pre-concentrate with WO3 grade of 54.32% and As grade of 5.41% was obtained. Finally, two stage magnetic cleaning were conducted on the second pre-concentrate for upgrading the second pre-concentrate, the mostly arsenic and gangues were removed from the second pre-concentrate. Finally, this novel processing achieved an effective recovery of fine wolframite from low grade tungsten ore with high arsenic. A concentrate with WO3 grade of 64.81% were obtained in new opened processing flowsheet, and As grade was less than of 0.5% in this concentrate. Furthermore, remarkable feature of this processing flowsheet is environment-friendly due to avoid using the toxic flotation reagents during the whole mineral processing.INTRODUCTIONTungsten is an essential metal and is the most widely used in industrial production such as national defense, electro-mechanical products, textile and leather sectors (Srivastava et al., 2000). Wolframite and scheelite are the main mineral sources of tungsten metal. With the continuous depletion of high-grade tungsten ores and the increasing demand for tungsten alloy, the low-grade refractory wolframite has become an important source for the tungsten industry. However, the large-scale utilization of low-grade refractory wolframite is disappointedly scarce, it is due to leanness of valuable minerals, complex mineral characterization and heavy use of toxic collectors."

Jan 1, 2016

By B. R. Maas

Processing material at a consistent rate through the continuous countercurrent decantation (CCD) thickeners presented a major bottle neck to unlocking additional value from one of the post-flotation streams at the Kansanshi processing plant. The manual operation of the thickeners making up the CCD train resulted in frequent imbalances and process upsets across units. The automation of CCD trains is notoriously tricky as the available manipulated variables are outnumbered by the parameters requiring regulation. The implementation of an advanced, auction-based, control philosophy to manage the movement of both slurry and solution inventory through the train addressed this parameter imbalance and lead to the stabilisation of the circuit. Since the commissioning of this control strategy: an additional 7 200 dmt/day (+36% throughput) of material can reliably be processed through the train, the number of monthly downtime events has reduced, and a saving of $354 000/month in reagent cost has been achieved. INTRODUCTION CCD operation Agitated leaching is a common method used in hydrometallurgy to extract a metal from its host ore. The host ore is initially finely ground to liberate the surfaces of the valued minerals and improve the kinetics of the leaching process. The ore is then contacted in an agitated tank with a lixiviant to dissolve the desired metal into the aqueous phase. The resultant slurry from the leaching process comprises of a solution pregnant in the metal salt and a barren ore. The first purification step taken after leaching is to wash the pregnant liquor solution (PLS) from the depleted ore. On the industrial scale, a set of between two to ten thickeners arranged in a countercurrent configuration is often used to carry out this separation (Perry & Chilton, 1973). The slurry is introduced at the head of the train and a wash solution at the tail. The two streams are pumped through each thickener countercurrent to one other. At each stage the residual PLS in the slurry is mixed and diluted down with the countercurrent overflow solution from the downstream thickener. The diluted slurry is fed into the thickener where the solids separate out through gravity sedimentation. The settled solids are channelled towards the underflow of the thickener by a rake mechanism and pumped downstream to the next thickener in the train. The outcome is the recovery of a PLS from the overflow of the lead thickener and a slurry of washed solids from the underflow of the tails thickener.

Jan 1, 2019

By Zhang Wei

"Pulang Copper Mine is located in Shangri-La county, Diqing Prefecture, Yunnan Province, with currently defined copper reserve of 4.8 million tonnes of copper metal and an average grade of 0.34% Cu. The mineralized zone in a dome shape is 2,300 m long, 600–800 m wide, and 1,000 m high. At present, the project is at a construction phase, with a first-stage mining and processing capacity of 12.5 million tonnes of ore per year. Ore haulage is planned via a drift and ore pass development system. The cavability of mineralized rock is classified into three types (II, III, IV), with type III accounting for more than 98% of the total reserves. The cavability of the orebody is generally considered to be good and block caving mining method is planned. Semi-autogenous grinding and ball milling are adopted for the mineral processing. The tailing dam construction method is a center-line method.INTRODUCTIONAs the China's largest underground nonferrous mine, Pulang Copper Mine is located in the northeast of Shangri-La County, Diqing Prefecture, Yunnan Province, China. It is 62 km northeast of Shangri-La city. Shangri-La County is a Tibetan-dominated, multiple-ethnic region, with a population of approximately 163,000 in 2009. The county is rich in mineral and water resources, thriving with tourism. Its economy is also dominated by agriculture. To date, the defined copper reserves at the Pulang Copper Mine is 4.8 million tonnes of copper metal at an average grade of 0.34% Cu. In addition, it contains a gold reserve of 145 tonnes, a silver reserve of 2,754 tonnes, a molybdenum reserve of 0.19 million tonnes, and a sulphur reserve of 10.5 million tonnes. The first-stage mining and processing capacity is planned at 12.5 million tonnes of ore per year. When it ramps up to this production rate, the mine will produce copper concentrate containing 50,000 tonnes of copper and molybdenum concentrate containing 968 tonnes of molybdenum. A drift and ore pass system is planned for ore haulage and mining method is a block caving method. Semi-autogenous grinding and ball milling are considered for mineral processing. A center line dam is adopted for the tailing dam construction (Feng, 2010; Feng, Li, & Wang, 2011; China Enfi Engineering Corporation, 2014)."

Jan 1, 2016

By T. C. King

AT the new Graham-Central Mill of Eagle-Picher, near Galena, Ill., material is simultaneously dewatered and introduced into the ball-mill scoop boxes by the use of variable-speed, 24-in. spiral classifiers. This arrangement has proved satisfactory in that the moisture content of the wet feed is reduced, thus allowing maintenance of density within the mills, and a constant feed at any desired rate to the mills is provided. In this mill the primary feed is wet crushed to minus ½ in. and preconcentrated by jigs to produce a bulk lead-zinc-iron concentrate and a waste tailing. The jig concentrate is then dewatered by a conventional dewatering classifier and stored in, bins ahead of the ball mills. Even though this material is well dewatered, the discharge at the bin bottom is sloppy because of the gradual drainage of the water from the surface of the particles. Conventional belt feeders with vertical cutoff gates are unsatisfactory because of the flushing characteristic of the wet sand.

Jan 1, 1950

By Joseph F. Shirley

A froth flotation plant to recover byproduct molybdenite must be designed specifically for the selected separation process and the type of ore being treated. However, some general rules concerning byproduct molybdenite plant design can be suggested that would be valid for most processes. The term byproduct molybdenite in this paper is limited to the recovery of molybdenite from copper porphyry ores, which essentially amounts to selectively separating molybdenite from copper and iron sulfide concentrates. In addition, most of the discussions would apply only to the almost universally used approach of floating the molybdenite and depressing the copper and iron sulfide minerals. Thickener It is considered essential that the feed to a byproduct molybdenite plant be thickened prior to being treated for molybdenite recovery for the following reasons: 1) Thickening allows the rougher and any preconditioning stages to be operated consistently at the optimum percent solids. 2) Generally, the size of equipment used for pretreatment and the rougher flotation is considerably smaller and less expensive than if the copper flotation concentrates were treated directly. 3) The removal of as much water as practical by thickening removes flotation reagents added in the copper flotation circuit, especially frother which can be quite troublesome in a byproduct molybdenite plant. Although thickening of the molybdenite plant feed has long been recognized as necessary to optimum metallurigcal results, very often this area has been designed poorly. The thickener should be sized so that it can produce a consistent thickened product containing 60 to 65% solids without the use of flocculants. Some plants are only able to thicken to 45 to 50% solids consistently while other plants let the percent solids vary from 45 to 60% over a 24-hr period. While the former is more desirable than the latter, in many cases better metallurgical results could be obtained if the percent solids could be held consistently close to 60%. Probably the simplest and most satisfactory system of handling the thickener underflow is to spigot it into a sump, using a sphincter type valve for the orifice, controlled by a density meter. Another satisfactory approach is to use an ODS type pump to meter the underflow. There are other methods of handling thickener underflow which work satisfactorily as long as the amount of copper concentrates entering the thickener is relatively uniform. However, they often fail to operate satisfactorily when there are large fluctuations in the thickener feed rate due to fluctuations in ore tonnage and/or copper grade. If the percent solids in the plant feed could be held close to 65%. There is the possibility the molybdenite plant tailing would not have to be thickened prior to the copper filter plant. This means that the molybdenite rougher float would have to be conducted at 45 to 50% solids with all the cleaner tailings returned to the rougher for dilution. Admittedly, in some situations this approach would be impossible, but if capital or plant area is limited, it should be considered.

Jan 1, 1981

By H. L. Talbot, A. J. Hall, R. S. Young

A method for the chemical differentiation of oxidized cobalt from sulphide cobalt in complex mine and concentrator products is of considerable interest to a number of mining companies throughout the world. At Rhokana Corporation, in Northern Rhodesia, for instance, cobalt is found chiefly in the sulphide form, as carrollite, but a cobalt oxide mineral also occurs irregularly underground, and relatively small quantities in the mill feed from time to time are difficult to detect visually at the concentrator. The oxide cobalt is not amenable to the sulphide flotation process used at Rhokana and nearly all of it passes into the tailing. The occasional period when high tailing losses occur with no appreciable change in the cobalt content of the mill feed has naturally been disturbing to the concentrator staff, who had no means of determining whether an influx of cobalt oxide ore or laxity on the part of the mill flotation operators as responsible for these results. A procedure for determining oxidized copper minerals, based on their solubility in dilute sulphuric acid saturated with SO2, has been used for many years in the copper industry. In the presence of cuprite the procedure fails, but otherwise it gives results quite satisfactory for control pur- poses. Empirical methods for determining oxidized forms of zinc and lead have been proposed, but they are generally unsatisfactory except in isolated cases of a very simple ore. No published procedure for determining oxidized forms of cobalt or siinilar elements has come to the authors'notice. Experimental Procedure There are obviously two alternatives in attacking a problem of this nature. One is to employ a solvent for oxide cobalt that will not dissolve cobalt sulphide. The other is to find a substance capable of oxidizing the cobalt sulphide to sulphate, which of course is water-soluble and can then be leached from the water-insoluble oxide cobalt. A representative sample of Rhokana cobalt oxide ore was secured, ground to minus 200-mesh, and carefully rolled to furnish a standard sample for this work. The analysis of this material gave the following percentages: Co, 0.55; Cu, 1.64; Fe, 2.98; S, 0.08; SiO2, 41.42; Al2o3, 11.20; CaO, 13.82; MgO, 3.80. As a source of cobalt sulphide a similar large sample of carrollite was prepared, analyzing Co, 23.36 per cent; Cu, 27.32; and S, 31.13. This sample of course was not pure carrollite, but was the best obtainable in quantities sufficient for a standard intended for considerable work. It was a pure sulphide sample, microscopic examination showing only carrollite, a little bornite, and gangue, with no visible sign of oxidized ore present. A wide variety of reagents as used in an effort to selectively dissolve oxide cobalt or to oxidize sulphide cobalt to

Jan 1, 1949

By G. S. P. Pereira, E. M. R. Santos

Acid mine drainage (AMD) is a serious environmental problem in mining activities. Prediction is the first step for evaluate the potential of acid generation and there are some possible methods to estimate it. They are described as static tests, which examine the balance between acid producing and acid consuming materials, and kinetic tests, which attempt to predict drainage quality overtime. Static methods are very simple and give a qualitative classification of the ores based on their potential to generate acid drainage. The characterization of ores to determine their acid producing potential is vital for the prevention of long term environmental liabilities. Laboratory tests were conducted under different methodologies to predict the acid producing potential of sulphide overburden, ore materials and tailings from a deposit of copper ore.

Jan 1, 1999

By A. J. S. Spearingt

The hydraulic transportation of relatively dense slurries for backfill purposes is typically carried through vertical and horizontal pipelines under gravity. Wear in backfill pipelines has become such a major problem that it is now one of the main reasons for the relatively slow build-up of placed backfill in spite of encouraging performance results in the gold-mining industry. This paper describes tests on pipeline wear that were carried out on classified tailings from the backfill plant at Vaal Reefs East. The testwork was carried out under various conditions of slurry relative density, slurry velocity, and pipeline diameter. Empirical relationships are presented for each transportation and pipeline parameter evaluated and, finally, overall relationships between pipeline conditions and transportation parameters are proposed.

Jan 1, 1992

By O. Kangal, G. Önal, G. Bulut, A. Gül

The recovery of copper and cobalt from the ancient copper slags, located in Küre-Turkey, is investigated. Fayalitic type of ancient copper slags containing 1.41% Cu, 0.53% Co and 53.16% Fe were subject to the experimental study. Magnetic separation, gravity concentration and flotation test were conducted. The results of magnetic separation and gravity concentration were not satisfactory. However, a concentrate containing 9.57% Cu could be produced with 44.7% recovery by flotation. Copper and cobalt that remained in the flotation tailings was recovered by roasting with sulfuric acid followed by hot water leaching. Complete cobalt dissolution was observed in 1 hour of roasting at 150°C and 3:1 acid/slag ratio, whereas the copper recovery was 83%.

Jan 1, 2004

By Peter Mason, Chris Walker, Mike Costello

The Phase I electrowinning system at KCGM's Fimiston Operation utilised the well proven technology of electrowinning precious metal values onto steel wool followed by calcining and direct smelting to bullion. The above process, while reliable and successful, was to some extent labour intensive. The Williams Mine in Ontario, Canada had published conditions under which precious metal values are electrowon onto woven stainless steel and directly removed using high pressure sprays. The current paper describes the testwork which was carried out to check the applicability of the Williams criteria to KCGM eluates arising from the elution of highly loaded carbon from the calcine circuit and lowly loaded carbon from the float tails circuit. Information is also given on the commissioning and opera­tional experience which occurred.

Jan 1, 1996

By S. H. Williston

IT IS not often that bore hole surveys can be checked by actual civil engineering methods. A recent Arizona survey was checked by normal surveying methods and the comparison of the results should be of value to both oil and mining men. During the summer of 1948 the Phelps-Dodge Corporation, at its Copper Queen property near Bisbee, Ariz., drilled a 1245 ft, 8 in. diam, churn drill-hole in a mineralized area and cased part of it, intending to use it to transfer mill tailings for stope fill. The hole, as frequently occurs, was not straight, and, in endeavoring to locate the bottom in the underground workings, they found no evidence of the hole at the underground coordinates directly below the surface location.

Jan 10, 1950

By J. C. Mostert, P. N. Roberts

Electric smelting at Rustenburg Platinum Mines Limited of nickel - copper concentrates containing platinum-group metals By J. C. MOSTERT Pr. Eng.. B.Sc. (Stellenboseh). M.Sc. Met. Eng. (Fellow) and P. N. ROBERTS Pr. Eng.. B.Sc. Eng. (Rand) (Fellow) The processes involved in the smelting in an electric furnace of nickel-copper concentrates containing platinumgroup metals are described. These are concentrate drying, pelletizing, pellet drying, and flux and pellets proportioning, followed by smelting of the concentrate in a submersible-electrode electric furnace, which is discussed in more detail. The furnace matte is blown in a converter to the white-metal stage and is cast in moulds, whilst the slag is granulated, ground, and floated for recovery of valuable metals and afterwards pumped to the tailing dam.

By J. S. Orphanos

The requirements for controlling air and stream pollution are a most timely subject for concerned people. The coal industry has taken great measures to reduce pollution. With the use of more efficient scrubbing equipment on their thermal dryers, pollution has been considerably reduced. Furthermore, eliminating thermal drying completely by using centrifugal dryers (Bird screen bowls) to dewater fine coal products and meet consumer specification has completely eliminated air pollution. The coal industry is taking steps to eliminate stream pollution and the hazards created by tailings ponds. These plants are closing their water circuit by recovering all solids in solid-bowl centrifugal equipment. In doing this effectively and economically the water discharged into the streams does not contain any refuse solids. This provides a safe economical method.

Jan 1, 1976

By W. E. Field

In the Coast Mountains approximately 110 miles north of Vancouver, the gold mine of Bralorne- Pioneer Mines Ltd. lies at an elevation of 3500 ft. The deepest or 41 level in the mine is at an elevation of 1400 ft below sea level, and the rock temperature on this level is 122.5°F. The thermal gradient is 1 "F increase in temperature for each 67.5 ft increase in depth. The mine produces between 500 and 600 tons per clay, utilizing an adit two internal shafts, and one service shaft from surface. The Queen shaft, extending from the 26 to the 41 level, IS the deepest, and most production now comes from this area. Stoping in by cut and fill using mill tailing hydraulic backfill.

Jan 12, 1963

By G. Gaidajis

Sea water samples collected from coastal waters in the vicinity of the Stratoni mining operations, in North-Aegean Sea were analysed for dissolved metals and elements. The maximum concentrations observed were one to three orders of magnitude lower than the environmental limit which coincides with the quality of sea waters suitable for swimming purposes. The quality of the coastal sea waters are affected by the natural mineralization of the area and the urban activities drained through local small rivers, the past mining activities present in the sea bottom sediments in the form of disposed tailings, and finally by the presence of the adjacent Mill-Flotation Plant. The relative importance of the treated final effluents discharged from the mining operations is considered to be insignificant.

Jan 1, 2006

By Mason P, Waller C

The Phase I electrowinning system at KCGM's Fimiston Operation utilised the well proven technology of electrowinning precious metal values onto steel wool followed by calcining and direct smelting to bullion. The above process, while reliable and successful, was to some extent labour intensive. The Williams Mine in Ontario, Canada had published conditions under which precious metal values are electrowon onto woven stainless steel and directly removed using high pressure sprays. The current paper describes the testwork which was carried out to check the applicability of the Williams criteria to KCGM eluates arising from the elution of highly loaded carbon from the calcine circuit and lowly loaded carbon from the float tails circuit. Information is also given on the commissioning and operational experience which occurred.

Jan 1, 1996

By R. J. Snow, S. N. Bhattacharya, D. D. Patil

Electrokinetics involves the movement of charged particles or water molecules in an applied dc field. Applications in particle fluid systems include the dewatering of soils, the densification of mill tailings and the treatment of coal wastes. In this paper, a novel application of this technique is discussed, wherein an electric field is applied to a coal-water slurry flowing through a pipe, resulting in lower pressure losses in the system. A unique experimental setup was designed and built to study the effect of an electric field (dc) on the wall shear stress for the flow of coal-water slurry through the pipe. The effect of solids concentration, current strength and pH on drag reduction across the system was quantified.

Jan 1, 2002

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

The Ardlethan tin concentrator had dete riorated considerably in metallurgical per- formance, maintenance and plant availability. This problem was compounded by low tin prices and International Tin Council quotas. Mid- 1976 figures were: throughput - 1300 tonnes per day, availability 83% and metal recovery 537. Major problems existed throughout the circuit from crushing to tailings disposal and with morale, staffing, and housekeeping. With the lifting of the quotas in June 1976, management provided funds to implement a major rehabilitation programme and some fifteen major projects, at a cost of about $1.3 million have been completed to date. The change in performance has been dram- atic, and end of 1977 performance figures are: throughput - 1700 tonnes per day, availability 90% and metal recovery 65%.

Jan 1, 1978