· · · Institutional Repository

The basis for this research project was that copper is lost in the Leaching Plant of the Nyrstar Budel zinc smelter. In the section of the Leaching Plant copper is lost, iron and aluminium are precipitated. Sampling of this process showed that the removal of copper from the solution is a function of pH and is mainly linked to the removal of aluminium. Tests with process fluids at stable pH levels showed that ferric ion and aluminium are mainly removed as jarosites at pH = 2 and pH =2,5. Copper is also removed at those pH values and is thus precipitated in jarosite form. This is the actual loss of copper. At pH = 3 and 3,5 more aluminium was removed, but now also in the form of aluminium hydroxides. It was found that copper is adsorbed onto these hydroxides and hence more copper is removed from the solution. At these pH values also jarosites are formed which permanently remove copper, but it is assumed that the adsorption onto aluminium hydroxides prevails. The result of the adsorption of copper onto these hydroxides is that it eventually returns to the FeP in which it has another chance to get lost to jarosites. Hence the only way to prevent copper from being lost is to prevent copper from reaching the FeP. One way to prevent copper from reaching the FeP is not using calcine to neutralise the acid from the SiHALO, but with alternative neutralising agents which do not contain copper. Another way is to prevent copper from leaving the SiHAL. Copper cementation is a method which can totally remove the copper dissolved in the SiHALO. Copper cementation with iron was found to yield the highest recovery and is thus also the most cost effective. Key words: Copper incorporation, jarosite, alunite, copper cementation

In this study metallurgical processing of two different kinds of zinc-bearing residues have been performed: Zinc A and Zinc B. These residues have been stored for over 15 years in Rotterdam Harbor. The chemical compositions of the residues have been determined and showed that zinc ferrite is a major phase present. Zinc ferrite is not soluble under normal alkaline and acidic conditions and is not recovered by the Waelz-process, which is commonly employed for such zinc-bearing residues. An innovative flowsheet for processing zinc ferrite-bearing residues has been developed during a pre-feasibility study, with goal to selectively recover zinc, including zinc from zinc ferrite. The innovative flowsheet consists of the following steps: (1) Water pre-washing: removing water soluble salts, in particular the chlorides in Zinc A (~9%). (2) 1st step alkaline leaching with caustic soda (NaOH): dissolving free ZnO into solution, for both water-washed Zinc A and original (unwashed) Zinc B. (3) Roasting of the first leach residue in the presence a suitable reagent: decomposing the zinc ferrite to free ZnO. (4) 2nd step alkaline leaching with NaOH: dissolving all free ZnO into solution. (5) Solution purification by cementation: removing impurities in particular lead and copper, by using zinc powder. (6) Electrowinning of zinc in NaOH solution: the purified zinc bearing solutions are subsequently precipitated to the final product of Zn metal. Optimal operating conditions for the processes are deduced from a literature review in which similar residues are processed. Additionally, optimal operating conditions for the conversion of zinc ferrite into zinc oxide has been investigated using synthetic zinc ferrite with addition of Mg(OH)2, Ca(OH)2, NaOH, or Na2CO3. Finally, Na2CO3 has been chosen as reagent and used in experiments with real zinc-bearing residues. Zinc A is water washed to remove the chlorides present. Then both the water washed residue of Zinc A, and Zinc B, are leached in an alkaline solution of 5M NaOH at 90˚C for 1 hour. Both zinc and lead are selectively extracted, leaving iron oxides and zinc ferrite in the residue. The filtercake is fused with Na2CO3 at 950˚C for 2 hours to convert zinc ferrite into zinc oxide. The calcined product is leached in fresh alkaline solution of 5M NaOH to recover zinc. The final residue is then water washed to remove residual sodium. The filtrates from the first and second leaching step are purified, with use of zinc dust, or directly used for electrowinning experiments. The removal efficiency of chloride, sodium and potassium during water washing of Zinc A were 62%, 41% and 71% respectively. Overall dissolution yields for Zinc A and Zinc B of zinc and lead were 82%, 80% and 64%, 78% respectively. Cementation of impurities (Pb, Cu, Cr) with zinc dust followed by an electrowinning step achieving a grade zinc deposit of 94%. Finally, it can be concluded that the combined hydro -and pyrometallurgical flowsheet is technically feasible. Furthermore, results can be improved further by optimization of major operating steps.

In current practice stope mines are designed using simple rules of thumb or using algorithms that do not find the true optimisation. Therefore a combinatorial optimisation algorithm was developed for optimising stope boundaries for sublevel stope mines. The profitability of a stope and the feasibility of a mine depends on the infrastructure that is needed to reach and operate the stope. An algorithm using Ant Colony Logic was developed for designing the access and longer sections of infrastructure. A third algorithm using hill climbing optimisation was developed for generating an optimised ore hauling system. For this thesis, the developed algorithms were used for generating an optimised design for the Kibali project underground sublevel stope mine. This design is then compared to the design by Randgold. It can be seen that the mining locations in both designs match. Because the optimisation algorithm takes the stope shapes into account a higher percentage of dilution is found.

This thesis studies the zinc extraction from spinels through hydro and pyrometallurgical processing. Two zinc-bearing spinels are covered: zinc ferrite ZnO·Fe2O3 to a limited extent and gahnite ZnO·Al2O3 as the main subject, compounds which are found naturally on the Earth’s crust as well as in industrial residues from the zinc industry, steel industry and others. Zinc ferrite contains 27 % of Zn and 33 % as ZnO; resource recovery from ferrite has been studied already in the past. On the other hand, the processing of gahnite, containing 35 % of Zn and 44 % as ZnO, is studied more extensively since research in the field of extractive metallurgy is effectively non-existent. Hence, the main objective of the present thesis is finding routes of treatment for this spinel. Zinc ferrite was produced synthetically at the CiTG/3mE labs by mixing equimolar amounts of ZnO and Fe2O3 at 1100 °C. Gahnite was produced by an analog method, a mixture of equimolar amounts of ZnO and Al2O3. The first approach was hydrometallurgical. Atmospheric hot acid leaching (4 M, 95 °C, 120 min, L/S 40) was performed with H2SO4, HCl and HNO3, resulting in a non-detected dissolution of the compound. Pressure leaching (90 min, L/S 40) was carried out in an autoclave with H2SO4 and HNO3, resulting in a low (2.9 %; 0.75 M, 140 °C, 3.6 bar) and a moderate extraction (22.2 %; 4.0 M, 250 °C, 39.7 bar) respectively. The second approach was pyrometallurgical processing (60 min dwell, 10 °C/min heating rate), divided into two sub-routes. A series of carbothermic tests (1:1.25 stoichiometric ratio) successfully led to a full reduction of the spinel at 1300 °C (99.90 % extraction of zinc). Aluminothermic tests (1.5:2 stoichiometric ratio) successfully resulted in a 99.98 % zinc extraction at 1200 °C. The mix of gahnite and ferrite with carbon at 1300 °C produced a 99.65 % extraction of the metal. Addition of ZnO to the previous mixture resulted in a 100 % extraction, at 1300 °C. Further experiments with gahnite at 1200 °C by adding SiO2, first with carbon and later with aluminium, resulted in a moderate 23.14 % and a low 4.69 % extraction correspondingly. Trials with CaO at 1400 °C created a glass residue and a slag, in each case. It is thus possible to establish the zinc extraction from gahnite ZnO·Al2O3 as follows: Route / Zinc extraction Atmospheric acidic leaching / Non-detected Pressure leaching / Low – Moderate Reduction with aluminium and silica / Low Reduction with carbon and silica / Moderate Carbothermic reduction / Full Aluminothermic reduction / Full

Vanadium is an important industrial metal that is mostly used as an alloying element in steel and to a lesser extent in titanium. It is generally produced in combination with other metals such as iron, titanium, and uranium. The typical production method begins with salt roasting of the ore or concentrate to produce a water soluble form of vanadium. Then the vanadium is leached, purified and precipitated as vanadium pentoxide. Most of the vanadium pentoxide produced is then combined with iron to create a ferrovandium alloy, via aluminothermic reduction, that is suitable for the iron and steel industry for alloying. Vanadium is also found in crude oil in a range that can vary from 10 to 1400ppm depending on the location. As the oil is processed and refined the vanadium becomes enriched. Specifically it is enriched in the fly ash of heavy fuel oil power plants. While the original concentration of vanadium in oil is very small the amount of oil extracted from the earth is very large and this makes fly ash a significant source of vanadium. The vanadium that is contained in the heavy oil fly ash can be recovered by two general methods: hydrometallurgical and pyrometallurgical. The hydrometallurgical route usually involves leaching, purification, and precipitation to produce V2O5. This V2O5 can then be used to produce ferrovanadium. The developmental pyrometallurgical route directly uses the fly ash in combination with a reductant (C, Al, or Si) and a source of iron to produce a ferrovanadium alloy. This thesis investigates one pyrometallurgical method, which uses two industrial waste products, heavy fuel oil fly ash and basic oxygen furnace dust, to produce a ferrovanadium alloy. The carbon contained within the fly ash is used as the reductant so only fluxing agents must be added to the charge. The result is a very efficient process that requires very few virgin raw materials. The results show that a ferrovanadium alloy with 15% vanadium can be produced by high temperature smelting of heavy oil fly ash and basic oxygen furnace dust. There are some impurities that remain in the metal, mainly nickel (~4%), sulfur (~5%), and carbon (~1.5%). The carbon to metal ratio had the largest effect on the final metal quality. The slag quality is also important to the final quality but more work must be done in this area. Simple water leaching tests were done on the as-received fly ash and the slag produced from smelting. The leach solution produced from the slag contained 100 times less metals than the solution produced from the as-received fly ash.

Kidd Mine has a production target of 2.5 million tonnes of ore per year in 2010. Seventy-five stopes are turned over to the next stope in order to achieve this annual production. Each turnover from stope to stope has an anticipated number of days based on the geomechanical relation between them. Due to the depth and size of the operation, it is crucial that this turnover takes place within the anticipated time to avoid delays in the mining sequence and cycle and set-backs in production. Currently delays in the stope turnover occur, this is called stope slippage. This thesis describes the occurrence and size of stope slippage in longhole mining, presents a system to identify and track the root causes of stope slippage and ranks the root causes of stope slippage at Kidd Mine. A flowchart was created to present the system of identifying and ranking root causes of stope slippage.

Raw magnesium chloride can be recovered using solution mining at a depth of a 1500 to 2000 meters. Underground caverns are formed in stacked layers of bischofite , carnallite and halite. The salt layers consist for a large part of NaCl and but contain a minimum of at least 35% magnesium chloride. Recovery is implemented by injection of fresh water in the salt layer, which becomes saturated with magnesium. The carnallite and bischofite layers largely consist of the less soluble NaCl, with interdispersed magnesium salts. A cylindrical model of magnesium recovery is presented that consists of a central open space, an annular space filled with a porous salt (NaCl) layer with an outer boundary that consists of a bischofite layer or carnallite layer. The central cavity is an open cylinder filled with a solution of NaCl and MgCl2. The presence of KCl can be disregarded. Fresh water is injected into the centre of the central cavity. Brine is extracted at a distance below the injection point. The natural convection flows of the fresh injection water in the cavity are investigated. Calculations show that the central cavity contains a solution of more or less constant composition except near the central axis of the cavity. Fresh injection water is lighter than the brine and therefore it migrates to the top of the cavity while mixing with the brine. The concentration of the brine near the injection point at the axis of the central cavity increases rapidly so that the roof will not be exposed to brines that are able to dissolve significant amounts of NaCl. Adjacent to the central cavity there is a concentric annular space, which consists of a skeleton of crystalline NaCl. This acts as a porous medium. The outer radius of the porous medium is adjacent to the undisturbed bischofite layer. At the outer radius the concentration of bischofite and sodium chloride are given by the saturated equilibrium conditions. Also at the outer radius there is a no flow boundary. At the inner radius the concentrations are constant due to the mixing conditions in the central cavity. The values of these concentrations are given by the cavity growth model. The Elder model is used to simulate natural convection flows through a porous annular cylinder of low permeability. The enhancement in the transfer rate due to natural convection flows with respect to mass diffusion is expressed in the Sherwood number. Simulations were conducted on porous media with a permeability of 2e-12 m^2 or less. The maximum enhancement factor resulting from the simulations is two and a half. For higher permeabilities the Brinkman model is used. The cavity growth model describes the rate of change of the inner and outer radius of the porous medium. The rate of change of the outer radius is determined by the diffusive mass flux of MgCl2. The rate of change of the inner radius is determined by the dissolution of the NaCl skeleton. However, the equilibrium concentration of NaCl in the presence of dissolved MgCl2 is rather small leading to a low dissolution rate of NaCl. Results of the cavity growth model are presented in terms of the inner and outer radius of the annular cylinder as a function of time. The cavity growth model requires enhancement factors in the order of 1e5 to get magnesium concentrations that are comparable to field data. Such high enhancement factors can possibly only be obtained using the Brinkman model.  

In order to meet the future demand of pellets for the blast furnace, the output of the grinding circuit of the Tata Steel IJmuiden pellet plant needs to be increased. As the power consumption of the grinding circuit is already at its maximum, the grinding itself and/or the grinding strategy needs to be optimised, leading to a higher circuit output while still achieving the required fineness. Further processes within the pellet plant require the grinding product to be of a certain fineness which is currently determined by measuring the Blaine number of the grinding circuit output, which is a measurement of the available surface within the grind. At the IJmuiden pellet plant, grinding is done in a closed grinding circuit, meaning that only particles smaller than a certain size are allowed to leave the circuit. Particles that are to coarse are screened out of the main flow of material and fed back to the grinding mills. The pellet feeds of the IJmuiden pellet plant are blends of different ores. The compositions of these blends are based on several factors such as availability, price and iron content. Properties that affect the grinding of the ores, such as initial fineness and grindability, however, are not taken into consideration in the selection of ores. Given this fact, a possibility to increase the grinding capacity might be found in treating the feed of the grinding circuit not as a single material, like is done at the moment, but as a collection of individual ores. This approach allows for a blending/grinding strategy to be designed based on overgrinding softer ores and leaving the harder ores coarser while still achieving the required overall fineness. As this would reduce the energy needed to achieve a certain fineness, this would increase the grinding capacity. Based on knowledge gained through a literature study on the binding mechanisms in iron ore green pellets, the strength and plasticity of green pellets is believed to be affected by such variations in individual ore fineness within the ore feed. To study the influence of individual ore fineness on green pellet strength and plasticity, an experimental study was designed that involved balling and testing of green pellets balled from four different feeds that only varied in individual ore fineness. In addition, the influence of strain rate on strength and plasticity of these same pellets was tested. It was found that variations in individual ore fineness had no significant influence on the strength and plasticity of green pellets. Pellet plasticity and strength were found to be strain rate dependent even for very low strain rates. Based on these experimental results, recommendations for further work and possible improvements in pelletising process monitoring were put forward.

After fifteen years of service, blast furnace #7 at IJmuiden’s Tata Steel operation was blown down on the 31st of August in 2006 for a small repair. The blow down and salamander tap were successfully completed and afterwards the furnace was quenched with water. All remaining liquids are solidified followed by excavation of the remaining skull. Copper was added to the last ore dump to distinguish the liquids prior to the quench. Around 300 holes were drilled in the skull, used for explosives. The cores were gathered and used for analysis. Several cores have been analyzed with X-ray fluorescence spectrometry; these rough data were the base of this study. Carbon lamellas were observed in certain areas of the skull. Their formation appears during slowly cooling of flowing hot metal. These lamellas confirm that part of the skull was solid previous to the blow down of blast furnace #7. Radial variation of silicon is not detected. Results do show a distinct boundary, which separates material with different concentrations of copper. This is possibly a result of early solidification of the skull. Confidential report Full version of the thesis can be requested at TATA Steel IJMUIDEN At the RD&T department reference source number: 153291

Los Bronces is an open pit mine located on a large porphyry copper-molybdenum deposit near Santiago, Chile. Annual production is around 240 000 tonnes of copper and 2 700 tonnes of molybdenum. The extracted ore has a relatively low average copper grade of around 0.6%. This results in relatively high processing costs and creates the need to search for solutions to reduce the costs for ore processing at Los Bronces. Sorting by sensor technology may be a solution to decrease these processing costs. Sensor based sorting is a technique where singular particles are mechanically separated on certain physical properties after determining these properties by a sensor or detector. Sensor based sorting is of relatively low cost compared to other concentration methods. It is an evolving technology that is reaching the requirements for implementation in ore processing operations. However, sensor based sorting is completely dependant on an appropriate sensor that allows distinguishing between sub-economic and economic ore material. The applicability of a near-infrared (NIR) spectroscopic sensor was investigated on a set of 150 mill pebbles from the Los Bronces processing plant to asses the feasibility of sensor based sorting. The NIR spectroscopic sensor allows characterisation of the reflected electromagnetic radiation from solids along the NIR wavelength regions of the electromagnetic spectrum (350-2500 nm). Mainly due to the vibrational effects of certain molecule bonds in these solids, absorption bands of lower reflected radiation are present around characteristic wavelength locations. These absorption bands can therefore be diagnostic for certain mineral presences in a solid. Copper minerals do not cause any characteristic absorption bands in the NIR region. Various hydrothermal alteration minerals that are associated with the formation of porphyry copper systems on the other hand do. These minerals include muscovite, illite, chlorite, tourmaline and kaolinite. Assessing the applicability of NIR spectroscopy for sensor based sorting of the Los Bronces mill pebbles was therefore focused on finding a relation between the copper grade and these alteration minerals. From general geological models it is known that copper mineralisation can be associated with certain zones of hydrothermal alteration. However, in practice the alteration zones and copper mineralisation usually form very complex systems due to several periods of intrusion, brecciation and overprinting of the pre-existing hydrothermal alterations. The test work showed that no direct correlation between the copper grade and a NIR spectral characteristic was present. However, it did prove to be possible to classify the mill pebbles based on the NIR spectral response and mineralogy determined from petrography and x-ray diffraction. This resulted in several groups of pebbles that proved to be identifiable by NIR spectroscopy and contained different average copper grades. One group of low grade pebbles was particularly interesting. This group contains 30% of all pebble samples with an average copper grade of 0.29%. This group can be easily identified with NIR spectroscopy by a high depth ratio between the 1900 and 2200 nm absorption feature and presence of an absorption feature around 2350 nm. These spectral features are caused by a low crystallinity of mica minerals and presence of chlorite respectively. It was shown that it is technically feasible to make some discrimination on the copper grade of the Los Bronces mill pebbles by the response from a NIR spectroscopic sensor. However, because the discrimination possibilities are limited to only one group of pebbles, a detailed economical analysis still has to prove the economical feasibility of NIR spectroscopy as a sensor sorting application for the Los Bronces mill pebbles. A preliminary economical analysis already showed that the economical feasibility is mainly dependant on the copper price and the processing costs of the ore.

This report introduces a method for the removal of zinc from aircraft aluminium-alloy scrap. The driving force for this research is the demand from the secondary aluminium industry for a low zinc content in the Al-alloy scrap, and the growing demand for aluminium in the world. Meanwhile large amounts of obsolete aircraft are stored because of the problems during recycling of the aluminium. The influence of coating on the aircraft scrap recycling is studied by treating the scrap in a de-coating process, regarding its influence on the melting process of the scrap. The coating accounts for approximately 1,6 wt% of the scrap. On a small scale, the melting of de-coated scrap shows much better results in coalescence and the separation of alloy from slag. On a larger scale, the melting results are also better for de-coated scrap, while the melting process of not de-coated scrap forms a reasonably coalescent alloy piece. The recyclability of aircraft scrap is studied in the presence of different salt fluxes. The addition of 10 wt% cryolite promotes the coalescence of the alloy and the separation of alloy and slag. However, this salt flux removes magnesium from the alloy. Addition of magnesium fluoride maintains or even increases the magnesium content in the alloy, but gives poor results in the melting process. The use of a higher salts-to-alloy ratio does not improve the melting results and possibly even counteracts the evaporation of zinc for both cryolite and magnesium fluoride as an addition. To improve the zinc removal from the aluminium alloy, a lance is used to blow argon gas into the alloy melt. The argon gas reduces the partial pressure of zinc, thereby promoting the evaporation of zinc from the melt. The tests are performed on an alloy with an initial zinc content of 2,42 %. A test with argon blowing lowers the zinc content to 1,88 %. However, a similar test without argon blowing results into an alloy with a zinc content of 2,11 %. The argon blowing only reduces the zinc content with 0,23 % compared to the similar test in which the same alloy is molten but no argon is blown into the melt.

The Baboto ore body is situated along the same geological structure as the Yalea and the Loulo3 ore bodies. It is located approximately 10km north east of the ROMPAD, 1km north of the village of Baboto. It stretches 5km to the north, a part of the Baboto North ore body lies in the Kofi permit. Relative little sampling has been done; core drilling every 25m, so all reserves are inferred. A block model has been made using ordinary kriging. This estimates the reserves at 11.9Mt with an average grade of 2.06 g/tonne. The ore body will be split up into two categories, all ore with a grade higher than 2.5 g/t shall be classified as high grade ore and all ore with a lower grade than 2.5 g/t shall be classified as low grade ore. The high grade ore will be processed at the processing plant. For the low grade ore the option of processing via a heap leaching process is being investigated. Metallurgical bottle roll tests have shown a possible plant recovery for the high grade ore of 89 to 94%. Heap leaching recovery is being estimated around 80%. Pit optimization using Gemcom’s Whittle software with a gold price of $1,200 per tonne produces 3 pits which would produce 5.74Mt of ore, or 462,166 ounces of gold content that will be processed. 301,164 ounces gold content is qualified as high grade ore, 161,002 ounces are from low grade ores. The main objective in mining is to have a continuous feed of the plant of 50,000 tonnes per month. This can best be done by first mining Baboto Center, than Baboto North after which Baboto South is the final area to be mined. This will result in a total mine life of 3.7 years. The total mining costs are assumed to be approximately $2.85 per tonne. The high grade ore will be transported to the plant by the same trucks as currently used for Gounkoto and 3 trucks shall be needed. The route from the Baboto ROM to the crusher at the ROMPAD will be approximately 12.5km, of which approximately 9km will be newly constructed. The low grade material from Gara and Yalea currently stockpiled next to the ROMPAD including the scats, are probably useful material for heap leaching. This is 1.57Mt at an average grade of 1.67 g/t. So the first 2.6 years the trucks will bring this material along on their way back to Baboto. This would only require a detour of 583m and 10 minutes of loading and dumping extra. The total transport costs for the high grade ore from Baboto to the crusher and the stockpile material form the ROMPAD to Baboto will be approximately $4.68 per tonne, including capital depreciation. The scats, just as the Baboto oxide ores, will not have to be crushed and can be used for heap leaching right away, the other non-oxide ores will have to be crushed. The leaching time for the ore on the heaps will be 60 days, after which it will be extracted via activated carbon and transported as high gold content liquid to the processing plant. Here it will undergo the same process as the comparable fluids at the plant. The estimated costs for heap leaching are $12.56 per tonne for the scats and $14.06 per tonne for all other ores. Whereas the processing costs for the high grade ore in the plant is $21.66 per tonne. 3 months will be needed for the entire infrastructure for the project to be constructed. With the G&A costs set at $5.98 per tonne of ore the total costs for all Baboto operations including the heap leaching of the ROMPAD stockpiles the costs will be $297.1M. This will produce 467,650 ounces of gold, which would give at a gold price of $1,200 per ounce a revenue of $561.2M. With an annual discount rate of 0% this would result in a NPV of $264.0M and an IRR at annual rate of 476%.

The Potgietersus Platinum mine is an open pit mine in the Bushveld Complex in South Africa. During the mining different rock types are mined, some of these are detrimental. The goal of the research done in this project is to assess the feasibility of Dual Energy X-ray Transmission (DE-XRT) to distinguish the detrimental ore types from the rest of the ore. There are two types of rock that contain the most platinum and Platinum Group Elements (PGE's): pyroxenite and pyroxenite B. The detrimental rock types are serpentinite, norite, calc- silicate and oxidized material. Detrimental rock types cause a lower recovery of the processing of the minerals. By removing the detrimental rock types the recovery could increase. DE-XRT is a sensor based sorting method based on density differences of the materials that have to be distinguished from each other. Samples of the different rock types have been scanned with a DE-XRT scanner and the pictures of these samples consist of pixels. Of each pixel the High Energy Absorption (HEA) and Low Energy Absorption (LEA) are known. With this data a color is allocated to each pixel. Ratios of these colors are plotted in graphs. From these graphs it can be concluded that a seperation can be made by using the ratio between the number of green and blue pixels. By using the green/blue ratio more than 60% of the calc silicate, 30% of the norite, 100% of the oxidized material and 60% of the serpentinite are removed. There is no loss of pyroxenite and pyroxenite B. It is recommended to do further research on other software or to improve the now used software. Also large scale experiments have to be conducted to find out if this method can be used in the mine itself. And a feasibility study has to be performed to see if a large enough percentage of the detrimental rock is removed to make the investment profitable.